Underground metal mine fan-shaped medium-length hole blasting anti-collision and vibration reduction collaborative control method

Abstract

The present application belongs to the technical field of underground metal mine blasting engineering, in particular to a kind of underground metal mine fan-shaped medium-length hole blasting anti-scouring-vibration reduction collaborative control method, with blasting energy control as core, three-step strategy, including the following steps: first, establish multi-index caving step optimization criterion system, optimize the caving step of multi-row hole blasting;Increase the delay time between the holes in the row;Second, the traditional fan-shaped upward medium-length hole distribution is divided into middle, two sides and the remaining three regions, first detonate the middle and two sides area holes;Finally, change the traditional single center drilling method to three drilling center method to arrange upward fan-shaped medium-length hole, establish "strength threshold-value-energy uniformity threshold-value" double constraint model, and propose a quantifiable drilling center distance determination method, to ensure that blasting energy release is more balanced and controllable, through the above steps, realize the fine regulation and collaborative release of blasting energy.

Description

Technical Field

[0001] This invention belongs to the technical field of underground metal mine blasting mining engineering, and in particular relates to a method for anti-impact and vibration reduction coordinated control of deep hole blasting in underground metal mines. Background Technology

[0002] Deep-hole fan-shaped blasting, as a highly efficient mining and rock breaking technology, is widely used in mining operations. Its hole pattern arrangement can fully utilize blasting energy and improve rock breaking efficiency. However, during the mining blasting process, due to unreasonable blasting parameter design or insufficient blasting free face, problems such as over- and under-excavation of blast holes can occur, leading to damage to roadway brow lines and subsequent blast holes, thus affecting blasting effectiveness, ore recovery rate, and mining safety. Excessive back impact can cause misalignment or blockage of subsequent blast holes, affecting their breaking effect and, in severe cases, even rendering them unusable, requiring re-drilling. Simultaneously, excessive back impact combined with vibration can cause widespread roadway brow line damage, increasing mining instability. When brow line damage is severe, the safety risks to charging personnel increase significantly. Therefore, during deep-hole fan-shaped blasting, measures should be taken to minimize back impact and blasting vibration to ensure the integrity and usability of subsequent blast holes, reduce brow line damage, and ensure the safety of workers.

[0003] The patents with publication numbers CN104818944A, CN120626169A, CN120556921A, CN113446006A, and CN102635356A involve aspects such as shock-proof and vibration-reduction control in deep-hole blasting of underground metal mines, and have achieved certain results in improving the directionality of blasting energy and enhancing the uniformity of ore and rock fragmentation. However, these technical solutions still have significant shortcomings, specifically: ① Patent CN104818944A involves the application of multi-center drilling, but it only limits the range of values ​​for the drilling center distance and does not provide a quantitative calculation or optimization method. This means that the solution only qualitatively proposes a multi-center drilling approach, but lacks corresponding methods for the specific implementation of the solution and how to adapt to different geological conditions. At the same time, the method does not consider the destructive effect of the blasting impact on the eyebrow line; ② CN120556921A involves the application of multi-free-face blasting technology. Although it utilizes the addition of voids in the fan-shaped boreholes to form a guiding effect, the arrangement of voids increases the amount of drilling work and occupies the charging space, affecting the release of blasting energy. At the same time, the guiding effect is greatly affected by geological conditions and blast source disturbances, and the stability is insufficient. Therefore, there are still significant limitations in improving blasting efficiency and controlling energy balance. ③ Although CN113446006A adopts a multi-free-face blasting structure formed by drilling tunnels and preparation wells, the free-face configuration is limited by the tunnel layout, making it difficult to flexibly control the direction of blasting energy. Furthermore, key parameters such as drilling center distance and timing delay are not optimized. Simultaneously, the construction of preparation wells is complex and involves a large workload, further reducing the applicability of this scheme in actual mining areas. ④ Both CN102635356A and CN120626169A employ multi-free-face deep-hole blasting methods, relying on cutting wells or tunnel structures to form free faces. However, their energy release paths are limited by spatial conditions, resulting in insufficient control over the blasting direction. Additionally, the construction of cutting wells involves a large workload and high breakthrough difficulty. The construction of free faces is highly dependent on structural stability; once the induction effect weakens due to fractured geological conditions, the risk of back-impact effect and eyebrow line damage remains difficult to effectively avoid. In summary, existing technical solutions still need improvement in quantitative design of drilling parameters, spatiotemporal distribution control of blasting energy, and the weakening of back-impact effect and prevention of eyebrow line stability. Summary of the Invention

[0004] To address the shortcomings of existing technologies, this invention provides a method for coordinated control of shock and vibration reduction in deep-hole blasting in underground metal mines. By quantitatively designing drilling parameters and controlling the spatiotemporal distribution of blasting energy based on the multi-free-face effect, it achieves refined regulation and coordinated release of blasting energy, reduces back impact and blasting vibration intensity, effectively prevents eyebrow line damage, and improves the safety and stability of underground metal mines. This method solves problems such as uneven energy distribution and significant impact damage during deep-hole blasting.

[0005] The technical solution of this invention is as follows:

[0006] A method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine includes the following steps:

[0007] Step 1: Based on the three indicators of over- and under-excavation volume, back-impact damage area, and energy transfer effect, determine the optimal blasting step distance for multi-row blasting. Within the optimal blasting step distance range, micro-delay detonation is adopted for the blasting holes between rows.

[0008] Step 2: Divide each row of blast holes into three areas: the middle, the two sides, and the remaining area. Use micro-delay blasting to detonate each area. The detonation sequence for each row of blast holes is: middle, two sides, and the remaining area blast holes.

[0009] Step 3: Set the drilling center of the deep hole in the fan shape to a multi-center layout, determine the optimal drilling center distance of multiple drilling centers, and arrange the number of blast holes according to the hole bottom distance.

[0010] Furthermore, the specific method for determining the optimal blasting step distance for multi-row blasting in step 1 is as follows:

[0011] Step 1-1: Under the condition of multi-row blasting, combined with the rock mass mechanical properties and blasting parameters, determine different caving step distance schemes, and set the failure criterion with the tensile strength of the rock mass as the threshold to determine whether the rock mass has reached the effective crushing condition.

[0012] Steps 1-2: Establish a numerical calculation model of the medium-deep hole stope and surrounding rock, and set monitoring points on the calculation model. For each blasting step scheme, perform simulation calculations on the numerical calculation model of the stope and surrounding rock, and record the changes in vibration velocity during the blasting process at each monitoring point to evaluate the transmission capacity and directionality of blasting energy; statistically analyze the over- and under-excavation of the blast hole face under different blasting step conditions; and observe the ore and rock damage range at 0m~0.5m behind the blast hole face to assess the area of ​​back-rush damage.

[0013] Steps 1-3: Compare the results of three evaluation indicators for each blasting step scheme: over- or under-excavation of the borehole face, back-impact damage area, and energy transfer effect. When the over- or under-excavation of the borehole face is the smallest, the back-impact damage area is the smallest, and the energy transfer to the forward free surface is the largest, the step scheme is determined to be the optimal blasting step scheme.

[0014] Steps 1-4: If the optimal blasting step distance schemes corresponding to the three evaluation indicators of over- or under-drilling amount of blast hole face, back-impact damage area and energy transfer effect are inconsistent, then the median value of the blasting step distance corresponding to the three is taken as the comprehensive optimal blasting step distance.

[0015] Furthermore, the specific method for determining the optimal drilling center distance among multiple drilling centers in step 3 is as follows:

[0016] Step 3-1: Establish numerical calculation models for different drilling center distances, set monitoring lines on the models, and perform simulation calculations on the models. Statistically calculate the equivalent stress between holes at the middle position of each borehole on the corresponding monitoring line when the drilling center distance is different, and calculate the mean and standard deviation of the equivalent stress between holes. Use the mean of the equivalent stress between holes to characterize the equivalent stress between holes, and use the standard deviation to characterize the energy uniformity coefficient of the degree of stress distribution non-uniformity.

[0017] Step 3-2: Fit the mean and standard deviation of the equivalent stress between holes under different drilling center distances obtained in Step 3-1 to obtain the functional relationship between the equivalent stress between holes, the energy uniformity coefficient and different drilling center distances.

[0018] Step 3-3: Based on the functional relationship between the equivalent stress between boreholes and different drilling center distances obtained in Step 3-2, the tensile strength of the rock mass is used as the critical value to obtain the drilling center distance strength threshold.

[0019] Step 3-4: Based on the functional relationship between the energy uniformity coefficient between blast holes and different drilling center distances in Step 3-1, when the drilling center distance is the drilling center distance intensity threshold, the energy uniformity coefficient of the blasting energy distribution reaches the natural equilibrium state is obtained, and it is denoted as the energy uniformity coefficient intensity threshold.

[0020] Step 3-5: Based on the energy uniformity coefficient intensity threshold obtained in Step 3-4, and analyzing the influence of rock mass quality, obtain the target threshold for the energy uniformity coefficient;

[0021] Step 3-6: Based on the functional relationship between the energy uniformity coefficient between boreholes and different drilling center distances in Step 3-1 and the target threshold of the energy uniformity coefficient in Step 5, the energy threshold of the drilling center distance is obtained.

[0022] Steps 3-7: Determine the optimal drilling center distance; take the minimum value of the drilling center distance intensity threshold and the drilling center distance energy threshold as the optimal drilling center distance.

[0023] Furthermore, in step 3-2, the functional relationships between the inter-hole equivalent stress and energy uniformity coefficient and different drilling center distances are as follows:

[0024] (1) The inter-hole equivalent stress between boreholes exhibits an exponentially decreasing functional relationship with different drilling center distances, and its expression is as follows:

[0025] σ e (S)=σ0e -kS (1);

[0026] Where: σ e(S) represents the functional relationship between the equivalent stress between boreholes and different drilling center distances; σ represents the equivalent stress between boreholes; σ0 represents the initial equivalent stress between boreholes, i.e., the equivalent stress between boreholes when S=0; k represents the energy attenuation coefficient; and S represents the drilling center distance.

[0027] (2) The functional relationship between the energy uniformity coefficient between boreholes and different drilling center distances is expressed as:

[0028] U(S)=Ae -mS (2);

[0029] In the formula: U(S) is the functional relationship between the energy uniformity coefficient between boreholes and different drilling center distances; U is the energy uniformity coefficient; A is the energy concentration coefficient, that is, the energy uniformity coefficient when S=0, which is the initial distribution state of blasting energy in the adjacent area of ​​the borehole; m is the energy diffusion index.

[0030] Furthermore, in step 3-3, when the equivalent stress between holes decays to the tensile strength of the rock mass, the expression for the strength threshold of the drilling center distance at this time is:

[0031] (3);

[0032] In the formula, S str σ is the rock drilling center-distance strength threshold. t This represents the tensile strength of the rock mass.

[0033] Furthermore, in steps 3-4, by combining equations (2) and (3), when the rock drilling center distance is the rock drilling center distance intensity threshold, the expression for the energy uniformity coefficient intensity threshold is obtained as follows:

[0034] (4);

[0035] In the formula, U th The intensity threshold is the energy uniformity coefficient.

[0036] Furthermore, in steps 3-5, the rock mass quality is introduced through the RMR value, and a rock mass quality correction factor is defined:

[0037] (5);

[0038] In the formula, ρ(RMR) is the rock mass quality correction factor;

[0039] At this point, the expression for the target threshold of the energy uniformity coefficient is:

[0040] U max (RMR) = ρ(RMR) × U th (6);

[0041] In the formula, U max(RMR) is the target threshold for the energy uniformity coefficient.

[0042] Furthermore, in steps 3-6, by combining equations (2) and (6), the expression for the energy threshold of the drilling center distance is obtained as follows:

[0043] (7);

[0044] In the formula, S u The energy threshold is the distance from the center of the rock drill.

[0045] Furthermore, in steps 3-7, the expression for the optimal drilling center distance is:

[0046] S*=min{S str , S u}(8;

[0047] In the formula, S * This is the optimal center distance for rock drilling.

[0048] Compared with the prior art, the beneficial effects of the present invention are as follows:

[0049] 1. The deep-hole blasting anti-impact and vibration reduction control method of the present invention can effectively reduce the back impact force of blasting, which can effectively address problems such as severe over-excavation and under-excavation, damage to the brow line, and severe post-blasting impact, thereby ensuring the integrity and usability of the rear row of blast holes and the brow line area, and thus improving the safety of the blasting process and the stability of the mining area.

[0050] 2. This invention proposes a quantifiable method for determining the drilling center distance by establishing a dual-constraint model of "strength threshold - energy uniformity threshold". This method uses the tensile strength of the rock mass as the mechanical criterion to derive the strength threshold formula, and establishes the energy uniformity threshold based on the energy attenuation law. Simultaneously, it introduces the rock mass quality index RMR as a correction factor for energy propagation and structural integrity, achieving adaptive optimization of the drilling center distance. Compared to existing schemes that only provide empirical intervals, this invention can quantitatively solve for the optimal center distance based on rock mass parameters and blasting energy characteristics, transforming drilling design from empirical to calculable, and exhibiting higher theoretical consistency, geological adaptability, and engineering stability.

[0051] 3. This invention employs a step-by-step micro-delay blasting strategy, proceeding from the center to the sides and then to the remaining areas. Combined with optimized control of the multi-row borehole blasting step distance and delay, a dynamic multi-free-face system is constructed within the borehole face. This ensures that blasting energy is primarily released towards the front and within the borehole, thus weakening the impact effect at the brow line and rear. This method simultaneously achieves a balance between improved energy utilization and damage control. Attached Figure Description

[0052] Figure 1Numerical calculation models for medium-deep hole mining areas and surrounding rock, and numerical calculation models for different drilling center distances, established for embodiments of the present invention;

[0053] Figure 2 This is a schematic diagram of the monitoring surface and monitoring points for the evaluation index of the ore collapse step distance optimization criterion in an embodiment of the present invention;

[0054] Figure 3 This is a diagram showing the arrangement of equivalent stress monitoring lines on the numerical calculation model for different rock drilling center distances in an embodiment of the present invention.

[0055] Figure 4 This is a schematic diagram of the partitioning of deep holes in each row of fan-shaped sections according to an embodiment of the present invention;

[0056] Figure 5 This is a schematic diagram of the blasting sequence within a fan-shaped blast hole array according to an embodiment of the present invention (from left to right in the diagram, the initial blast hole, the blast hole in the middle of the detonation, and the blast holes on both sides of the detonation).

[0057] Figure 6 This is a schematic diagram illustrating the principle of generating multiple free surfaces through internal blasting in a fan-shaped borehole array according to an embodiment of the present invention.

[0058] Figure 7 This is a schematic diagram of the multi-row hole blasting and ore-breaking step distance scheme according to an embodiment of the present invention;

[0059] Figure 8 This is a schematic diagram of the arrangement of the three central boreholes in an embodiment of the present invention;

[0060] Figure 9 This is a schematic diagram showing the location of the three rock-drilling centers according to an embodiment of the present invention;

[0061] In the attached diagram: 1. Fan-shaped medium-deep hole; 2. Blast hole; 3. Equivalent stress monitoring line; 4. Blast hole in the central area; 5. Blast holes in the two side areas; 6. Blast holes in the remaining area; 7. Free surface; 8. Drilling center point. Detailed Implementation

[0062] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be further described in detail below with reference to the accompanying drawings and embodiments. It should be understood that the specific embodiments described herein are merely illustrative of the invention and do not limit its scope.

[0063] The dimensions of the medium-deep hole stope structure involved in this embodiment are approximately 16.5m × 13m. The blasting for the medium-deep hole stope mining adopts an upward fan-shaped arrangement of medium-deep holes 1, with a diameter of Φ64mm. In the blasting parameter design, the minimum resistance line and row spacing are both taken as 1.5m, and the hole bottom distance is taken as 1.8m. The blasting filling adopts an intermittent filling structure, with filling section lengths of 1.0m and 3.0m respectively, and the detonation method is orifice detonation.

[0064] like Figures 1-9 As shown, a method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine includes the following:

[0065] Step 1: Based on the three indicators of over- and under-excavation volume, back impact damage area and energy transfer effect, determine the optimal blasting step distance for multi-row blasting. Within the optimal blasting step distance range, micro-differential detonation is adopted for the blasting holes between rows to achieve balanced release of blasting energy, weakening of back impact effect and improvement of stope stability.

[0066] Step 1-1: Under multi-row blasting conditions, combining the rock mass mechanical properties and blasting parameters, and using a resistance line of 1.5 m as a benchmark, establish blasting step distances of 1.5 m, 3.0 m, 4.5 m, 6.0 m, and 7.5 m, respectively. Figure 7 As shown, a failure criterion was set with a rock mass tensile strength of 4.58 MPa as the threshold to determine whether the rock mass has reached the effective fracture condition.

[0067] Steps 1-2: Establish a numerical calculation model of the medium-deep hole stope and surrounding rock. Set monitoring points at the stope boundary locations of the numerical calculation model of the medium-deep hole stope and surrounding rock, such as... Figure 2 As shown, for each blasting step scheme, the numerical calculation model of the stope and surrounding rock is simulated, and the vibration velocity change during the blasting process at each monitoring point is recorded to evaluate the energy transfer capability and directionality of the blasting; the over- and under-excavation of the borehole face under different blasting step conditions is statistically analyzed; and the ore and rock damage range is observed at 0m~0.5m behind the borehole face to assess the area of ​​back impact damage.

[0068] In this embodiment, LSDYNA is used to simulate and analyze the blasting process and energy characteristics. A numerical calculation model of the medium-deep hole stope and surrounding rock is established in this embodiment. Figure 1 As shown, the overall model dimensions are 30m × 25m × 5~12m, and the modeling unit is m–kg–s. A fluid-structure interaction method is introduced during the calculation to describe the interaction between the explosive detonation products and the rock-mineral medium. To ensure computational accuracy, the element mesh size is controlled within the range of 0.1~0.2m. The HJC constitutive model is used for the rock mass in the numerical calculation. The basic physical and mechanical parameters of the rock include: density ρ0 = 2700 kg / m³, compressive strength σ... c =82.79MPa, tensile strength σ t =4.58MPa, elastic modulus E=60.33GPa, the explosive material adopts the MAT_HIGH_EXPLOSIVE_BURN high-performance explosive material model, and its detonation process is described by the JWL equation of state; the air medium adopts the MAT_NULL empty matter material model, and the linear polynomial equation of state EOS_LINEAR_POLYNOMIAL is used.

[0069] In this embodiment, the numerical calculation model for medium-deep hole mining and surrounding rock is selected as shown in the attached figure. Figure 2 The A-A borehole profile shown is used as a representative section for the statistical comparison of over- and under-excavation; the profile 0.5m behind the B-B borehole profile is used as a comparative section for evaluating the degree of impact damage after blasting; at the same time, monitoring point C is set at the top of the boundary of the free face stope in the sector-shaped deep hole to record the change of vibration velocity during blasting, so as to evaluate the ability of blasting energy to be transferred to the free face.

[0070] Steps 1-3: Compare the results of three evaluation indicators for each blasting step scheme: over- and under-excavation of the borehole face, back-impact damage area, and energy transfer effect. The optimal blasting step scheme is determined when the over- and under-excavation of the borehole face is minimized, the back-impact damage area is minimized, and the energy transfer to the forward free surface is maximized. According to simulation results, the over- and under-excavation of the borehole face is minimized when the blasting step is 3.0m, at which point the over- and under-excavation of the borehole face is 13.87m. 2 When the ore-breaking step distance is 4.5m, the back-rush damage area is the smallest, at which point the back-rush damage area is 19.38m². 2 When the ore-breaking step distance is 4.5m, the kinetic energy transfer at the monitoring point is the maximum, and the particle vibration velocity is 2.54m / s at this time.

[0071] Steps 1-4: If the optimal caving step distance schemes corresponding to the above three evaluation indicators are inconsistent, the median value of the caving step distances corresponding to the three shall be taken as the comprehensive optimal step distance parameter. In this embodiment, the comprehensive optimal step distance parameter is 4.5m.

[0072] Steps 1-5: Based on the optimal blasting step distance of 4.5m, increase the micro-delay initiation time of the blast holes between rows to reduce the "obstruction" effect of the front row of blast holes on the blasting of the rear row of blast holes, thereby effectively reducing the back impact force of multi-row blasting.

[0073] Step 2: As Figures 4-6 As shown, each row of blast holes is divided into three areas: the middle, the two sides, and the remaining area. Each area is detonated by micro-differential blasting. The detonation sequence of each row of blast holes is the middle, the two sides, and the remaining area. By detonating the blast holes in the middle and the two sides first, multiple free surfaces are created within the row of blast holes, which induces the blast to act forward and within the row.

[0074] Step 2-1: In the blasting of each row of fan-shaped blast holes within the optimal ore-breaking step range, firstly, use micro-differential blasting to preferentially detonate the blast holes 4 in the central area, increasing the initial free surface within the row; then detonate the blast holes 5 in the side areas, continuing to increase the initial free surface 7 within the row, forming multiple free surfaces 7 within the blast hole row surface, such as... Figures 5-6 As shown; in this embodiment, one blast hole is set in the central area 4, and two blast holes are set in the two side areas 5.

[0075] Step 2-2: After the blasting of the boreholes in the central and lateral areas is completed, the remaining boreholes in area 6 are detonated using a micro-delay detonation method. At this time, the blasting energy is transmitted forward and into the blasting zone along the existing free surface. The crushed rock is mainly thrown to the front and sides of the stope, significantly reducing the reflection and impact of blasting gas in the eyebrow direction and reducing the backlash. Figure 6 As shown.

[0076] Step 3: Set the drilling center of the deep hole 1 in the fan shape to a multi-center hole layout, determine the optimal drilling center distance of multiple drilling centers, and arrange the number of blast holes according to the hole bottom distance.

[0077] In this embodiment, the traditional single-center drilling method is changed to a three-center drilling method. Specifically, blast holes are arranged vertically upwards at the centerline of the tunnel, and a drilling center is equidistantly positioned on both sides perpendicular to the centerline, resulting in three drilling center points 8. The number of blast holes 2 is designed based on a hole bottom distance of 1.8m. Figure 8 , Figure 9 As shown;

[0078] Step 3-1: Establish numerical calculation models for different rock drilling center distances, such as... Figure 1 As shown, to calculate the equivalent stress between holes at different drilling center distances, an equivalent stress monitoring line 3 is set 1 m away from the eyebrow line in the calculation model. The eyebrow line refers to the boundary line between the square roadway and the medium-deep hole stope in the calculation model, as shown in the attached figure. Figure 3 As shown in Table 1, when the optimal blasting step distance is 4.5m, the equivalent stress between holes at the middle position of each blast hole on the corresponding equivalent stress monitoring line 3 is statistically analyzed when different drilling center distances are used. The mean and standard deviation of the equivalent stress between holes are calculated. The mean of the equivalent stress between holes is used as the equivalent stress between holes, and the standard deviation is used as the energy uniformity coefficient characterizing the degree of uneven stress distribution.

[0079] Table 1 Values ​​of Equivalent Stress and Energy Uniformity Coefficients Between Holes at Different Drill Center Distances

[0080] ;

[0081] Step 3-2: Fit the mean and standard deviation of the equivalent stress between holes under different drilling center distances obtained in Step 3-1 (see Table 1) to obtain the functional relationship between the equivalent stress between holes, the energy uniformity coefficient and different drilling center distances.

[0082] (1) Functional relationship between the equivalent stress σ between boreholes and different drilling center distances S e (S) exhibits exponential decay, and its expression is:

[0083] σ e (S)=10.902e -0.708S (1);

[0084] (2) Establish the functional relationship U(S) between the energy uniformity coefficient U between blast holes and different drilling center distances S. U represents the degree of energy concentration between holes. When U is large, it indicates that the blasting energy concentration is high and the stress distribution is uneven. Its variation with the drilling center distance can be expressed as follows:

[0085] U(S) = 3.038e -0.939S (2);

[0086] Step 3-3: In the process of determining the center distance of rock drilling, in order to ensure that the rock fracture zone formed between adjacent blast holes can be interconnected, a strength threshold model is established with the tensile strength of rock mass as the critical criterion. Based on the functional relationship between the equivalent stress between blast holes and different center distances of rock drilling in Step 3-2, the strength threshold of the center distance of rock drilling is obtained.

[0087] When the equivalent stress between the holes decays to the tensile strength σ of the rock mass t When the strength is 4.58 MPa, the corresponding rock drilling center distance strength threshold S is... str Therefore, we can conclude that:

[0088] (3);

[0089] Step 3-4: Based on the functional relationship U(S) between the energy uniformity coefficient U between the boreholes and different drilling center distances S obtained in Step 3-1, when the drilling center distance is the drilling center distance strength threshold, that is, when the equivalent stress between the boreholes just decays to the tensile strength of the rock mass, the energy uniformity coefficient of the blasting energy distribution reaches the natural equilibrium state is obtained, and it is denoted as the energy uniformity coefficient strength threshold.

[0090] Combining equations (2) and (3), when the equivalent stress between holes just decays to the tensile strength of the rock mass, we obtain the energy uniformity coefficient strength threshold U. th :

[0091] (4);

[0092] Step 3-5: Based on the energy uniformity coefficient intensity threshold U obtained in Step 3-4 th The influence of rock mass quality was analyzed, and the target threshold for energy uniformity coefficient was obtained.

[0093] Rock mass quality is introduced through RMR=59, where RMR (rock mass rating) is the geomechanical classification of rock mass, used to reflect the influence of structural plane development on energy propagation and stress attenuation. A rock mass quality correction factor is defined as follows:

[0094] (5);

[0095] At this point, the target threshold U of the energy uniformity coefficientmax (RMR) is:

[0096] (6);

[0097] Step 3-6: Based on the functional relationship U(S) between the energy uniformity coefficient U between boreholes and different drilling center distances S in Step 3-1, and the target threshold U of the energy uniformity coefficient in Step 3-5. max (RMR), combining equations (2) and (6), we obtain the rock drilling center distance energy threshold S. u ;

[0098] (7);

[0099] Steps 3-7: Determine the optimal drilling center distance. Take the minimum value of the drilling center distance intensity threshold and the drilling center distance energy threshold as the optimal drilling center distance S. * ;

[0100] S*=min{1.225m, 0.944m}=0.944m (8).

[0101] This invention discloses a fan-shaped deep-hole blasting anti-impact and vibration reduction coordinated control method for underground metal mines. Centered on blasting energy control, the method employs a three-step strategy, including the following steps: First, establishing a multi-index blasting step distance optimization criterion system to optimize the blasting step distance for multi-row blasting; increasing the micro-difference time between rows of blast holes to reduce the "obstruction" effect of the front row of blast holes on the blasting of the rear row, thereby reducing the back impact force of multi-row blasting; Second, dividing the traditional fan-shaped upward-facing deep-hole distribution into three regions: the middle, the two sides, and the remaining area. By first detonating the blast holes in the middle and the two sides, the blasting energy is reduced... Multiple free surfaces are created within the borehole face to induce blasting action forward and within the borehole. Finally, the traditional single-center drilling method is changed to a three-center drilling arrangement with upward fan-shaped medium-deep holes. A dual-constraint model of "intensity threshold - energy uniformity threshold" is established, and a quantifiable method for determining the drilling center distance is proposed to ensure more balanced and controllable blasting energy release. This further provides a parameter basis for anti-impact and vibration reduction coordinated control. In this embodiment, after three steps of anti-impact and vibration reduction coordinated control, the over-excavation and under-excavation area is reduced to 12.36 m², the back-impact damage area is reduced to 11.92 m², and the eyebrow line damage area is 0.03 m². 2 This significantly improved blasting effectiveness and mining stability.

Claims

1. A method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine, characterized in that, Includes the following steps: Step 1: Based on the three indicators of over- and under-excavation volume, back-impact damage area, and energy transfer effect, determine the optimal blasting step distance for multi-row blasting. Within the optimal blasting step distance range, micro-delay detonation is adopted for the blasting holes between rows. Step 2: Divide each row of blast holes into three areas: the middle, the two sides, and the remaining area. Use micro-delay blasting to detonate each area. The detonation sequence for each row of blast holes is: middle, two sides, and the remaining area blast holes. Step 3: Set the drilling center of the deep hole in the fan shape to a multi-center hole layout, determine the optimal drilling center distance of multiple drilling centers, and arrange the number of blast holes according to the hole bottom distance. The specific method for determining the optimal drill center distance among multiple drill centers is as follows: Step 3-1: Establish numerical calculation models for different drilling center distances, set monitoring lines on the models, and perform simulation calculations on the models. Statistically calculate the equivalent stress between holes at the middle position of each borehole on the corresponding monitoring line when the drilling center distance is different, and calculate the mean and standard deviation of the equivalent stress between holes. Use the mean of the equivalent stress between holes to characterize the equivalent stress between holes, and use the standard deviation to characterize the energy uniformity coefficient of the degree of stress distribution non-uniformity. Step 3-2: Fit the mean and standard deviation of the equivalent stress between boreholes under different drilling center distances obtained in Step 3-1 to obtain the functional relationships between the equivalent stress between boreholes, the energy uniformity coefficient, and different drilling center distances, as follows: (1) The inter-hole equivalent stress between boreholes exhibits an exponentially decreasing functional relationship with different drilling center distances, and its expression is as follows: s e (S)=σ0e -kS (1); Where: σ e (S) represents the functional relationship between the equivalent stress between boreholes and different drilling center distances; σ represents the equivalent stress between boreholes; σ0 represents the initial equivalent stress between boreholes, i.e., the equivalent stress between boreholes when S=0; k represents the energy attenuation coefficient; and S represents the drilling center distance. (2) The functional relationship between the energy uniformity coefficient between boreholes and different drilling center distances is expressed as: U(S)=Yes -mS (2); In the formula: U(S) is the functional relationship between the energy uniformity coefficient between boreholes and different drilling center distances; U is the energy uniformity coefficient; A is the energy concentration coefficient, that is, the energy uniformity coefficient when S=0, which is the initial distribution state of blasting energy in the borehole adjacent area; m is the energy diffusion index; Step 3-3: Based on the functional relationship between the equivalent stress between boreholes and different drilling center distances obtained in Step 3-2, and using the tensile strength of the rock mass as the critical value, the drilling center distance strength threshold is obtained, expressed as: （3）； In the formula, S str σ is the rock drilling center-distance strength threshold. t Tensile strength of the rock mass; Step 3-4: Based on the functional relationship between the energy uniformity coefficient between blast holes and different drilling center distances in Step 3-1, when the drilling center distance is the drilling center distance intensity threshold, the energy uniformity coefficient of the blasting energy distribution reaches the natural equilibrium state is obtained, and it is denoted as the energy uniformity coefficient intensity threshold. Step 3-5: Based on the energy uniformity coefficient intensity threshold obtained in Step 3-4, and analyzing the influence of rock mass quality, obtain the target threshold for the energy uniformity coefficient; Step 3-6: Based on the functional relationship between the energy uniformity coefficient between boreholes and different drilling center distances in Step 3-1 and the target threshold of the energy uniformity coefficient in Step 3-5, the energy threshold of the drilling center distance is obtained. Steps 3-7: Determine the optimal drilling center distance; take the minimum value of the drilling center distance intensity threshold and the drilling center distance energy threshold as the optimal drilling center distance.

2. The method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine, as described in claim 1, is characterized in that... The specific method for determining the optimal blasting step distance for multi-row blasting in step 1 is as follows: Step 1-1: Under the condition of multi-row blasting, combined with the rock mass mechanical properties and blasting parameters, determine different caving step distance schemes, and set the failure criterion with the tensile strength of the rock mass as the threshold to determine whether the rock mass has reached the effective crushing condition. Steps 1-2: Establish a numerical calculation model of the medium-deep hole stope and surrounding rock, and set monitoring points on the calculation model. For each blasting step scheme, perform simulation calculations on the numerical calculation model of the stope and surrounding rock, and record the changes in vibration velocity during the blasting process at each monitoring point to evaluate the transmission capacity and directionality of blasting energy; statistically analyze the over- and under-excavation of the blast hole face under different blasting step conditions; and observe the ore and rock damage range at 0m~0.5m behind the blast hole face to assess the area of ​​back-rush damage. Steps 1-3: Compare the results of three evaluation indicators for each blasting step scheme: over- or under-excavation of the borehole face, back-impact damage area, and energy transfer effect. When the over- or under-excavation of the borehole face is the smallest, the back-impact damage area is the smallest, and the energy transfer to the forward free surface is the largest, the step scheme is determined to be the optimal blasting step scheme. Steps 1-4: If the optimal blasting step distance schemes corresponding to the three evaluation indicators of over- or under-drilling amount of blast hole face, back-impact damage area and energy transfer effect are inconsistent, then the median value of the blasting step distance corresponding to the three is taken as the comprehensive optimal blasting step distance.

3. The method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine, as described in claim 1, is characterized in that... In steps 3-4, combining equations (2) and (3), when the rock drilling center distance is the rock drilling center distance intensity threshold, the expression for the energy uniformity coefficient intensity threshold is obtained as follows: （4）； In the formula, U th The intensity threshold is the energy uniformity coefficient.

4. The method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine, as described in claim 3, is characterized in that... In steps 3-5, the rock mass quality is introduced through the RMR value, and a rock mass quality correction factor is defined: （5）； In the formula, ρ(RMR) is the rock mass quality correction factor; At this point, the expression for the target threshold of the energy uniformity coefficient is: U max (RMR) = ρ (RMR) × U th (6); In the formula, U max (RMR) is the target threshold for the energy uniformity coefficient.

5. The method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine, as described in claim 4, is characterized in that... In steps 3-6, combining equations (2) and (6) yields the expression for the energy threshold of the drilling center distance: （7）； In the formula, S u The energy threshold is the distance from the center of the rock drill.

6. The method for coordinated control of shock absorption and vibration reduction in deep-hole blasting in a sector-shaped underground metal mine, as described in claim 5, is characterized in that... In steps 3-7, the expression for the optimal drilling center distance is: S*=min{S str , S u }(8); In the formula, S * This is the optimal center distance for rock drilling.

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