Rock roadway full-face deep-hole slotting blasting method

Abstract

This application discloses a method for full-section deep-hole cut-out blasting in rock tunnels, relating to the field of blasting technology. To reduce the confinement effect of surrounding rock, the blasting method includes the following steps: setting a main cut-out hole in the cut-out area; uniformly arranging multiple auxiliary blasting holes along the boundary line of a first rhombus centered on the main cut-out hole; and uniformly arranging multiple secondary cut-out holes along the boundary lines of the second, third, and fourth rhombuses; placing explosives in each blast hole; and then detonating them according to a specified detonation sequence to generate superimposed stress and tensile stress on the rock. This application is applicable to blasting scenarios in rock tunnel excavation.

Description

Technical Field

[0001] This application relates to the field of blasting technology, and in particular to a method for full-section deep-hole slotting blasting in rock tunnels. Background Technology

[0002] As my country's mineral resource development extends to deeper areas, deep mine tunneling faces severe challenges such as high ground stress and complex geological conditions. Traditional tunneling blasting technology is affected by the confinement effect of surrounding rock in complex deep environments. The stress waves generated by shallow-hole blasting attenuate rapidly during propagation, significantly reducing rock-breaking efficiency. A large amount of blasting energy is absorbed by the surrounding rock and fails to be effectively converted into breaking work. Under the high ground stress conditions at depth, the constraint force of the surrounding rock increases dramatically, making it difficult for traditional blasting technology to overcome the confinement effect. As a result, the overall construction efficiency cannot meet the time and quality requirements of modern mine construction. Summary of the Invention

[0003] In view of this, the present application provides a method for full-section deep-hole slotting blasting in rock shafts, which facilitates the reduction of the surrounding rock confinement effect.

[0004] This application provides a method for full-section deep-hole blasting in rock tunnels, comprising the following steps: setting a main blasting hole in the blasting area; evenly distributing multiple auxiliary blasting holes along the boundary line of a first rhombus centered on the main blasting hole; outside the first rhombus, sequentially setting a second, third, and fourth rhombus centered on the main blasting hole from the main blasting hole outwards; and evenly distributing multiple secondary blasting holes along the boundary lines of the second, third, and fourth rhombuses, the secondary blasting holes having the same length as the main blasting hole; placing explosives in the main blasting hole, secondary blasting holes, and auxiliary blasting holes; wherein, the main blasting hole, from bottom to top, sequentially includes: an initial kinetic energy explosive section and a main plugging mud section; the secondary blasting holes, from bottom to top, sequentially include: a high-pressure throwing explosive section at the bottom of the hole, an isolation mud section, and a hole-mouth crushing and unloading section. The system consists of a pressure explosive section and a secondary plugging mud section; a projectile explosive section is installed at the bottom of the ejection aid hole; the detonation proceeds in the following order: the initial kinetic energy explosive section of the main slotting hole and the orifice breaking and pressure relief explosive sections of all secondary slotting holes on the boundary line of the second rhombus detonate simultaneously; the projectile explosive sections in all ejection aid holes on the boundary line of the first rhombus detonate simultaneously; the high-pressure projectile explosive sections at the bottom of all secondary slotting holes on the boundary line of the second rhombus detonate simultaneously and the orifice breaking and pressure relief explosive sections of all secondary slotting holes on the boundary line of the third rhombus detonate simultaneously and the orifice breaking and pressure relief explosive sections of all secondary slotting holes on the boundary line of the fourth rhombus detonate simultaneously; and the high-pressure projectile explosive sections at the bottom of all secondary slotting holes on the boundary line of the fourth rhombus detonate simultaneously.

[0005] Optionally, the lengths of the auxiliary throwing holes are all greater than the length of the main cut hole, and the propelling explosive section in the auxiliary throwing holes is arranged in the bottom extended section of the blast hole.

[0006] Optionally, the length of the high-pressure throwing explosive section at the bottom of the secondary cut hole: L1 = (0.35 - 0.4) × Hc; the length of the isolation stemming section of the secondary cut hole: L 泥 = (0.1 - 0.12) × Hc; the length of the hole-opening fragmentation and pressure-relieving explosive section of the secondary cut hole: L2 = (0.25 - 0.3) × Hc; the length of the secondary stemming section of the secondary cut hole: L 堵 = Hc - L1 - L 泥 - L2; where Hc is the length of the secondary cut hole.

[0007] Optionally, according to the basic quality index BQ of the rock mass, adjust the lengths of each section of explosive. Specifically: when BQ > 550, multiply the length of each section of explosive by 1.15 as the final length of each section of explosive; when 450 < BQ ≤ 550, multiply the length of each section of explosive by 1.1 as the final length of each section of explosive; when 350 < BQ ≤ 450, multiply the length of each section of explosive by 1.0 as the final length of each section of explosive; when 250 < BQ ≤ 350, multiply the length of each section of explosive by 0.9 as the final length of each section of explosive; when 150 < BQ ≤ 250, multiply the length of each section of explosive by 0.8 as the final length of each section of explosive; when BQ ≤ 150, multiply the length of each section of explosive by 0.75 as the final length of each section of explosive.

[0008] Optionally, the charge density of the initial kinetic energy explosive section of the main cut hole is 1.15 - 1.35 kg / m; the detonator of the main cut hole is arranged at the bottom of the main cut hole; the charge density of the high-pressure throwing explosive section at the bottom of the secondary cut hole is 1.15 - 1.35 kg / m, and the charge density of the hole-opening fragmentation and pressure-relieving explosive section is 0.95 - 1.15 kg / m; detonators are arranged at the bottoms of both the high-pressure throwing explosive section and the hole-opening fragmentation and pressure-relieving explosive section of the secondary cut hole; the initiation delay between the high-pressure throwing explosive section and the hole-opening fragmentation and pressure-relieving explosive section at the bottom of the hole is 40 ms.

[0009] Optionally, the blasting method further includes the step of grading and setting the spacing between the cut holes according to the lithological conditions: when the rock mass quality index RQD ≥ 85% or the number of joint sets Jn ≤ 2 sets, the cut hole spacing S is calculated according to the formula: ,

[0010] for calculation, with a value range of 900 - 1120 mm; when 60% ≤ RQD < 85% or 2 < Jn ≤ 3 sets, the cut hole spacing S is calculated according to the formula: ,

[0011] Calculations are performed, with a value range of 680~900mm; when RQD<60% or Jn>3 groups, the slot hole spacing S is calculated according to the formula: ,

[0012] Calculations were performed, with values ​​ranging from 400 to 680 mm; where RQD is the rock mass quality index, Jn is the number of joint groups, and S is the spacing between the cut holes. The diameter of the borehole. For the detonation velocity of the explosive, For the density of the explosive, It represents the uniaxial compressive strength of the rock.

[0013] Optionally, the blasting method further includes the step of: determining the spacing S of the cut holes according to the rock strength coefficient f or the joint spacing: when the rock strength coefficient f ≥ 12 or the joint spacing > 1.5m is a complete rock mass area, the spacing S of the cut holes ranges from 400 to 600 mm; when the rock strength coefficient 8 ≤ f < 12 or the joint spacing 0.5 to 1.5m is a moderately complete rock mass area, the spacing S of the cut holes ranges from 600 to 800 mm; when the rock strength coefficient f < 8 or the joint spacing < 0.5m is a fractured rock mass area, the spacing S of the cut holes ranges from 800 to 1120 mm; where f is the rock strength coefficient and S is the spacing between the cut holes.

[0014] Optionally, the diameters of the main slotting hole, the blasting aid hole, and the secondary slotting hole are determined based on rock parameters; specifically, the diameter of the main slotting hole is determined according to the formula:

[0015] ,

[0016] Perform calculations;

[0017] The diameter of the blasting aid hole is determined by the formula: ,

[0018] Perform calculations;

[0019] The diameter of the inner layer cut hole is determined by the formula: ,

[0020] Perform calculations;

[0021] The diameter of the outer layer slot hole is determined by the formula: ,

[0022] Perform the calculations; where D1, D2, D3, and D4 are the diameters of the main slotting hole, the auxiliary blasting hole, the inner layer slotting hole, and the outer layer slotting hole, respectively. It represents the uniaxial compressive strength of the rock. ρ is the rock density; RQD is the rock mass quality index, ranging from 0 to 100%; f is the rock firmness coefficient; Jn is the joint group number; K5 is the main slotting hole depth correction coefficient, ranging from 1.15 to 1.25; K6 is the auxiliary blasting hole compensation correction coefficient, ranging from 0.9 to 1.1; K7 is the inner layer slotting hole breakage correction coefficient, ranging from 0.95 to 1.05; K8 is the outer layer slotting hole transmission correction coefficient, ranging from 0.9 to 1.0; the inner layer slotting hole is the secondary slotting hole on the boundary of the second and third rhombuses; the outer layer slotting hole is the secondary slotting hole on the boundary of the fourth rhombus.

[0023] Optionally, the blasting method further includes: the first rhombus having a throwing aid hole at least at each vertex, and the second, third, and fourth rhombuses having a secondary slotting hole at least at each vertex; the vertex of the first rhombus coinciding with the midpoint of the side of the second rhombus; the vertex of the second rhombus coinciding with the midpoint of the side of the third rhombus; and the vertex of the third rhombus coinciding with the midpoint of the side of the fourth rhombus.

[0024] This application provides a method for full-section deep-hole slotting blasting in rock tunnels. Multiple auxiliary slotting holes and secondary slotting holes are set along a rhomboid boundary line centered on the main slotting hole. Different explosive distribution and detonation sequences are used in the main slotting hole, auxiliary slotting holes, and secondary slotting holes. The explosive stress wave generated by the explosion in the main slotting hole, combined with the phase reversal effect generated by the explosive stress wave in the auxiliary slotting holes distributed in the rhomboid configuration, produces a combined effect of tensile and compressive stress on the rock. Simultaneously, the projectile explosive segments in all auxiliary slotting holes along the first rhomboid boundary line detonate, further increasing the tensile and compressive stress on the rock between the auxiliary slotting holes at the bottom of the deep hole. This also ejects previously generated rock fragments from the blast hole, providing free space for subsequent explosions, weakening the overall rock integrity, and reducing the confinement effect of the surrounding rock in the deep hole. The initial kinetic energy explosive segment of the main slotting hole and the orifice crushing and decompression blasting of all secondary slotting holes along the second rhomboid boundary line further contribute to the blasting process. When all the explosive sections are detonated simultaneously, and the high-pressure thrown explosive sections at the bottom of all the secondary slotting holes on the boundary line of the second rhombus and the breaking and pressure-relieving explosive sections at the openings of all the secondary slotting holes on the boundary line of the third rhombus are detonated simultaneously, and the high-pressure thrown explosive sections at the bottom of all the secondary slotting holes on the boundary line of the third rhombus and the breaking and pressure-relieving explosive sections at the openings of all the secondary slotting holes on the boundary line of the fourth rhombus are detonated simultaneously, explosive stress waves are generated between each of the simultaneously detonated holes in directions that are parallel to the axial direction of the main slotting hole, perpendicular to the axial direction of the main slotting hole, and diagonally cross in space, thus generating superimposed tensile and compressive stresses in the surrounding rock in multiple directions. At the same time, in space, the breaking and pressure-relieving explosive sections at the openings of the secondary slotting holes on the boundary lines of the second, third, and fourth rhombuses are detonated first, followed by the high-pressure thrown explosive sections at the bottoms. The impact force generated by the later detonation has a throwing effect on the upper layer of broken rock generated by the earlier detonation, providing free space for subsequent blasting and making it easier to reduce the confinement effect of the surrounding rock. Attached Figure Description

[0025] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are only some embodiments of this application. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0026] Figure 1 This is a schematic diagram of the process of a full-section deep-hole blasting method for rock shafts according to an embodiment of this application;

[0027] Figure 2 This is a schematic diagram of the arrangement of the main cut holes, auxiliary blasting holes and secondary cut holes in a full-section deep-hole blasting method for rock shafts according to an embodiment of this application.

[0028] Figure 3This is a schematic diagram of the distribution of explosives in the main cut hole, auxiliary blasting hole, and secondary cut hole of a full-section 5m deep hole cut blasting method for rock tunnel excavation according to an embodiment of this application.

[0029] Figure 4 This is a schematic diagram of the blast hole layout for a full-section 5m deep-hole slotting blasting method for rock shaft excavation according to an embodiment of this application;

[0030] In the diagram: 1. Cutting area; 11. Main cutting hole; 111. Initial kinetic energy explosive section; 112. Main plugging mud section; 12. First rhombus; 13. Second rhombus; 14. Third rhombus; 15. Fourth rhombus; 16. Auxiliary throwing hole; 161. Projectile explosive section; 17. Secondary cutting hole; 171. Bottom high-pressure projectile explosive section; 172. Isolation mud section; 173. Orifice breaking and pressure relief explosive section; 174. Secondary plugging mud section; 2. Detonator; 3. Collapse zone; 4. Peripheral hole; 5. Bottom hole. Detailed Implementation

[0031] The embodiments of this application will now be described in detail with reference to the accompanying drawings.

[0032] It should be understood that the described embodiments are merely some, not all, of the embodiments in this application. All other embodiments obtained by those skilled in the art based on the embodiments in this application without inventive effort are within the scope of protection of this application.

[0033] Example 1

[0034] See Figure 1 and Figure 2 This application provides a method for full-section deep-hole blasting in rock shafts, comprising the following steps:

[0035] S100. A main slotting hole 11 is provided in the slotting area 1;

[0036] S101. A plurality of blasting aid holes 16 are evenly distributed on the boundary line of the first rhombus 12 centered on the main slotting hole 11.

[0037] S102. On the outside of the first rhombus 12, from the main slot hole 11 outward, a second rhombus 13, a third rhombus 14 and a fourth rhombus 15 are provided in sequence with the main slot hole 11 as the center; and a plurality of secondary slot holes 17 are evenly distributed on the boundary lines of the second rhombus 13, the third rhombus 14 and the fourth rhombus 15, and the secondary slot holes 17 have the same length as the main slot hole 11.

[0038] S103. Explosives are placed in the main slotting hole 11, the secondary slotting hole 17, and the ejection aid hole 16; wherein, the main slotting hole 11 is provided with, from bottom to top, an initial kinetic energy explosive section 111 and a main plugging mud section 112; the secondary slotting hole 17 is provided with, from bottom to top, a bottom high-pressure ejection explosive section 171, an isolation mud section 172, a hole opening crushing and pressure relief explosive section 173, and a secondary plugging mud section 174; and an ejection explosive section 161 is provided at the bottom of the ejection aid hole 16.

[0039] S104. The detonation proceeds in the following order: the initial kinetic energy explosive section 111 of the main slotting hole 11 and the orifice breaking and pressure relief explosive section 173 of all secondary slotting holes 17 on the boundary line of the second rhombus 13 are detonated simultaneously; the projectile explosive section 161 in all projectile aid holes 16 on the boundary line of the first rhombus 12 is detonated simultaneously; the high-pressure projectile explosive section 171 at the bottom of all secondary slotting holes 17 on the boundary line of the second rhombus 13 and the orifice breaking and pressure relief explosive section 173 of all secondary slotting holes 17 on the boundary line of the third rhombus 14 are detonated simultaneously; the high-pressure projectile explosive section 171 at the bottom of all secondary slotting holes 17 on the boundary line of the third rhombus 14 and the orifice breaking and pressure relief explosive section 173 of all secondary slotting holes 17 on the boundary line of the fourth rhombus 15 are detonated simultaneously; and the high-pressure projectile explosive section 171 at the bottom of all secondary slotting holes 17 on the boundary line of the fourth rhombus 15 is detonated simultaneously.

[0040] In this embodiment, the initial kinetic energy explosive segment 111 in the main slotting hole 11 generates the initial cracks and rock fragmentation and movement space for deep-hole slotting blasting; the explosive stress wave generated in the main slotting hole 11 produces a phase reversal effect on the wall surface of the auxiliary throwing hole 16 on the boundary line of the first rhombus, and then forms tensile stress in the rock in the rhombus configuration space. It works together with the subsequent compressive stress wave to break the rock through the combined action of tensile stress and compressive stress, promote rock fragmentation, and form the initial space, providing sufficient free surface and compensation space for the subsequent blasting of the auxiliary throwing hole 16 and the secondary slotting hole 17.

[0041] Multiple ejection holes 16 are evenly distributed along the boundary line of the first rhombus 12. Explosive sections 161 are placed at the bottom of each ejection hole 16. Each ejection hole 16 can hold a certain amount of rock debris. When all the explosive sections 161 in the ejection holes 16 along the boundary line of the first rhombus detonate simultaneously, they generate further tensile and compressive stress on the rock between the ejection holes 16 at the bottom of the deep hole. Simultaneously, the ejection action of the explosive sections 161 throws the previously generated rock debris out of the ejection holes 16, providing an initial free surface for subsequent explosions and weakening the overall rock structure. This reduces the confinement effect of the surrounding rock in deep holes, concentrates the explosive stress waves of the main cut hole 11 and the secondary cut hole 17, especially the explosive stress waves of the cut holes symmetrical along the auxiliary blast hole 16, reduces the squeezing effect between the whole rocks, improves the effect of the explosive pressure on the rocks to crush the rocks, guides the expansion of rock cracks, and improves the explosive effect of the cut zone 1; it throws the rock debris in the auxiliary blast hole 16 out of the auxiliary blast hole 16, reduces the manual handling of the rock debris in the blast hole, improves the tunneling efficiency, and at the same time provides sufficient free surface and compensation space for the subsequent blasting of the secondary cut hole 17.

[0042] The first rhombus 12, the second rhombus 13, the third rhombus 14, and the fourth rhombus 15 are all centered on the main cut hole, forming a regular spatial structure between the main cut hole 11 and the blast holes on the boundary lines of each rhombus. This ensures that when each blast hole is detonated, a uniform combined effect of tensile and compressive stress is generated between the main cut hole, the ejection aid holes 16 on the boundary lines of each rhombus, and the secondary cut holes 17. Multiple secondary cut holes 17 are distributed along the boundary lines of the second rhombus 13, the third rhombus 14, and the fourth rhombus 15. The secondary cut holes 17 utilize the free surface formed by the explosion of the initial kinetic energy explosive segment 111 in the main cut hole 11, and detonate sequentially according to the above detonation sequence, gradually expanding the volume of the cut area 1 cavity. The secondary cut holes 17 have the same length as the main cut hole 11, forming full-section deep holes in the rock. These deep holes constitute geometric weak surfaces, separating the surrounding rock mass to be excavated from the retained rock mass in the geometric weak surfaces formed by the deep holes, forming a blasting cutting line.

[0043] Under deep-hole conditions, the rock clamping effect at the bottom of the hole is strong, requiring sufficient energy in the lower section of the borehole to overcome the clamping. The explosive stress wave generated by the initial kinetic energy explosive section 111 in the main cut hole 11 breaks the rock in the deep hole. The shock wave generated by the initial kinetic energy explosive section 111, combined with the explosive gas and the blocking effect of the main plugging mud section 112, instantaneously compresses the rock in the main cut hole 11, causing the rock to crack and / or break. The main plugging mud section 112 is pushed out of the borehole by the force of the shock wave and the explosive gas. After the hole is opened, the gas pressure in the main cut hole 11 drops instantly, causing the rock in the main cut hole 11 to contract. This causes the rock in the main cut hole 11 to generate more cracks, further expand the original cracks, and / or break the rock into fragments, destroying the integrity of the deep hole rock. The initial kinetic energy explosive section 111 generates a shock wave, and the residual force of the explosive gas ejects the fragments generated in the main cut hole 11 out of the main cut hole, providing sufficient free surface and compensation space for the subsequent blasting of the auxiliary blasting hole 16 and the secondary cut hole 17.

[0044] The initial kinetic energy explosive section 111 of the main cut hole 11 and the orifice breaking and pressure relief explosive section 173 of all secondary cut holes 17 on the boundary line of the second rhombus 13 are detonated simultaneously; the high-pressure throwing explosive section 171 at the bottom of all secondary cut holes 17 on the boundary line of the second rhombus 13 and the orifice breaking and pressure relief explosive section 173 at the bottom of all secondary cut holes 17 on the boundary line of the third rhombus 14 are detonated simultaneously; and the high-pressure throwing explosive section 171 at the bottom of all secondary cut holes 17 on the boundary line of the third rhombus 14 and the orifice breaking and pressure relief explosive section 173 at the bottom of all secondary cut holes 17 on the boundary line of the fourth rhombus 15 are detonated simultaneously. When 73 are detonated simultaneously, explosive stress waves are generated between each detonated borehole, propagating in multiple directions: parallel to the axial direction of the main cut hole 11, perpendicular to the axial direction of the main cut hole 11, and diagonally intersecting in space. This results in superimposed tensile and compressive stresses on the surrounding rock in multiple directions. While there may be a certain axial angle deviation between the main cut hole 11 and the secondary cut hole 17 during actual construction, their axes are essentially parallel. Therefore, for ease of description, directions parallel to or perpendicular to the axial direction of the main cut hole 11, the blasting aid hole 16, and / or the secondary cut hole 17 are all described. The direction is parallel or perpendicular to the axial direction of the main cut hole 11; the superimposed tensile and compressive stresses generated on the surrounding rock by the above-mentioned multiple boreholes in multiple directions include at least: the direction of the line connecting the initial kinetic energy explosive section 111 of the main cut hole 11 and the explosion position of the orifice crushing and pressure relief explosive section 173 of the secondary cut hole 17 on the second rhombus 13; the direction of the line connecting the explosion position of the high-pressure throwing explosive section 171 at the bottom of the secondary cut hole 17 of the second rhombus 13 and the explosion position of the orifice crushing and pressure relief explosive section 173 of the secondary cut hole 17 of the third rhombus 14; and the direction of the high-pressure throwing explosive section 171 at the bottom of the secondary cut hole 17 of the third rhombus 14. The direction of the line connecting the explosive section 171 and the explosion position of the secondary slotting hole 17 of the fourth rhombus 15, and the direction of the line connecting the explosion positions of the secondary slotting holes 17 that are simultaneously detonated on the second rhombus 13, the third rhombus 14, and the fourth rhombus 15; each explosive section will generate at least an explosive stress wave in a direction parallel to and / or perpendicular to the axial direction of the main slotting hole 11 and in the above-mentioned connecting directions when it explodes; at the same time, the explosive stress waves generated in various directions interact with each other, producing a superimposed tensile stress and compressive stress on the rock, thereby promoting rock fragmentation.

[0045] The initial kinetic energy explosive section 111 of the main slotting hole 11 and the boundary line of the second rhombus 13, the orifice breaking and pressure relief explosive section 173 of all secondary slotting holes 17 on the boundary line of the second rhombus 13, the orifice breaking and pressure relief explosive section 173 of all secondary slotting holes 17 on the boundary line of the second rhombus 13 and the boundary line of the third rhombus 14, and the boundary line of the fourth rhombus 15 of all secondary slotting holes 17 on the boundary line of the third rhombus 14. All the secondary cut holes 17 on the line have their orifice crushing and decompression explosive sections 173 detonated in a diagonally intersecting manner in space and simultaneously in time. During the intersecting detonation, the stress waves generated by the explosions of adjacent cut holes propagate in the rock mass and have a superposition effect, producing superimposed stress. When two or more stress waves meet in the same area, according to the principle of wave superposition, each stress wave maintains its independent propagation characteristics. The total stress at the intersection point is the vector sum of the individual stress waves, resulting in a tensile effect on the rock at the intersection point. The cross-detonation generates tensile stress, enhancing the rock fragmentation effect. Furthermore, the spatial cross-detonation causes oblique cracks in the rock, and the subsequent explosive stress waves shear these cracks, promoting further rock fragmentation near the cracks. The spatial cross-detonation, occurring simultaneously in time, facilitates the generation of superimposed compressive and tensile stresses on the rock in the axial direction parallel to the main cut hole 11 between the two simultaneously detonated explosive sections, as well as superimposed compressive and tensile stresses in the axial direction perpendicular to the main cut hole 11 between the two simultaneously detonated explosive sections. The explosive stress wave generated by the initial detonation undergoes phase reversal at the wall of the auxiliary ejection hole 16, forming tensile stress in the rock within the rhomboid configuration space. Since the auxiliary ejection hole 16 and secondary cut hole 17 are substantially or completely parallel to the main cut hole 11 in the axial direction, the cross-detonated explosive positions generate superimposed compressive stresses on the rock parallel to the borehole axis and tensile stresses perpendicular to the borehole axis, improving the rock fragmentation effect and reducing the surrounding rock confinement effect.

[0046] In each secondary cut hole 17 on the boundary lines of the second, third, and fourth rhombuses, the orifice-breaking and pressure-relief explosive section 173 detonates first, reducing the constraint of the rock on the secondary blockage mud section 174 at the orifice-breaking hole 17. This blasts the rock at the orifice opening, creating a space larger than the orifice diameter. This provides a free surface and compensation space for the blasting of the high-pressure throwing explosive section 171 at the bottom of the hole, creating a larger rock fragmentation throwing channel space. This reduces the explosion energy consumption of the high-pressure throwing explosive section 171 at the bottom of the hole, lowering the explosive consumption per unit. Furthermore, the orifice-breaking and pressure-relief explosive section 173 detonates first, creating a rock fragmentation containing space at the orifice opening. The impact force generated by the subsequently detonated high-pressure throwing explosive section 171 at the bottom of the hole throws the upper layer of rock fragmentation generated by the first detonation, which helps to overcome the clamping effect of the surrounding rock under deep hole conditions.

[0047] The high-pressure throwing explosive sections at the bottom of all the secondary slots on the boundary line of the fourth rhombus are detonated in sequence. This further fractures and / or breaks the rock in the slotted area 1 caused by the previous blasting, and throws the generated rock fragments out of the blast holes, providing a free surface and compensation space for the extension of rock fissures and the movement of broken rocks during subsequent blasting.

[0048] In traditional slotting methods, frequent short-cycle operations increase the time cost of drilling, charging, and blasting processes, and also exacerbate equipment wear and personnel fatigue. However, the full-section deep-hole slotting blasting method for rock tunnels proposed in this application can adapt to complex geological conditions at depth, improve single-cycle advance, significantly increase borehole utilization, reduce explosive consumption, and also facilitate overcoming the clamping effect of surrounding rock under deep-hole conditions.

[0049] In some embodiments, 0-1 ms after controlled detonation, the orifice breaking and pressure relief explosive sections 173 of all secondary slotting holes 17 on the boundary line between the initial kinetic energy explosive section 111 of the main slotting hole 11 and the second rhombus 13 are detonated simultaneously. The geometric interference of the explosive stress wave generated by the explosion of the orifice breaking and pressure relief explosive sections 173 of the secondary slotting holes 17 on the boundary line between the initial kinetic energy explosive section 111 and the second rhombus 13 with the rhombus spatial configuration stress wave propagation between them and the wall of the assisted throwing hole 16 is used to improve the energy concentration of the explosive stress wave.

[0050] After a 5ms delay, the projectile explosive segments 161 in all the auxiliary projectile holes 16 on the boundary line of the first rhombus 12 detonate simultaneously, forming a synergistic rock-breaking effect with the explosive stress wave generated by the explosion of the explosive in the main slot hole 11 and the explosive gas. 30ms after controlled detonation, the high-pressure projectile explosive segments 171 at the bottom of all the secondary slot holes 17 on the boundary line of the second rhombus 13 detonate simultaneously, generating a rock-shattering effect on the secondary slot holes 17 above the boundary line of the second rhombus 13, with the bottom of the blast hole pointing downwards and the opening pointing upwards. At the same time, it impacts the rock around the secondary slot holes 17 on the boundary line of the third rhombus 14, causing it to loosen and creating rock fissures in the oblique direction. This facilitates the shearing effect between the explosive stress wave of the detonation explosive segment 173 at the opening of the secondary slot holes 17 on the boundary line of the third rhombus 14 and the rock fissures in the oblique direction, improving the rock-shattering effect and reducing the surrounding rock confinement effect.

[0051] After a 5ms delay, the pressure-relieving explosive sections 173 at the openings of all the secondary slotting holes 17 on the boundary line of the third rhombus 14 detonate simultaneously, creating a delayed detonation. This facilitates the rapid stretching effect of the rock between the two explosion stress waves formed by the cross-detonation of the explosions of the high-pressure thrown explosive sections 171 at the bottom of all the secondary slotting holes 17 on the boundary line of the second rhombus 13 and the pressure-relieving explosive sections 173 at the openings of all the secondary slotting holes 17 on the boundary line of the third rhombus 14, thereby improving the rock-breaking effect and reducing the surrounding rock confinement effect.

[0052] Sixty milliseconds after controlled detonation, the high-pressure thrown explosive sections 171 at the bottom of all the secondary slotted holes 17 on the boundary line of the third rhombus 14 detonate simultaneously. After a delay of 5 milliseconds, the breaking and decompression explosive sections 173 at the openings of all the secondary slotted holes 17 on the boundary line of the fourth rhombus 15 detonate simultaneously. The time-delayed detonation and spatial cross-detonation between the high-pressure thrown explosive sections 171 at the bottom of all the secondary slotted holes 17 on the boundary line of the second rhombus 13 and the breaking and decompression explosive sections 173 at the openings of all the secondary slotted holes 17 on the boundary line of the third rhombus 14 produce a similar explosion effect.

[0053] Ninety milliseconds after controlled detonation, the high-pressure throwing explosive sections 171 at the bottom of all secondary cut holes 17 on the boundary line of the fourth rhombus 15 detonate simultaneously. This not only breaks up the rock but also throws loose rocks above, improving the explosion effect of the cut zone 1 and increasing tunneling efficiency. By precisely controlling the detonation delay, stress waves from different cut holes arrive in phase in the target fracture area, achieving constructive superposition and significantly enhancing the stress intensity in the area. At the same time, the geometric symmetry of the rhombus hole arrangement ensures the consistency of the stress wave propagation path, which is conducive to the formation of stress concentration in the area between the cut holes at the center of the rhombus and the vertices of the rhombus, thus improving the delayed fracture rate.

[0054] In some embodiments, 30ms after controlled detonation, the orifice breaking and depressurization explosive sections 173 of all the secondary slots 17 on the boundary line of the third rhombus 14 detonate simultaneously, shattering the rocks at the orifice openings and creating a space larger than the diameter of the orifice, while also loosening the rocks around the secondary slots 17 on the boundary line of the third rhombus 14.

[0055] After a 5ms delay, the high-pressure throwing explosive sections 171 at the bottom of all the secondary cut holes 17 on the boundary line of the second rhombus 13 detonate simultaneously. The explosions of the high-pressure throwing explosive sections 171 at the bottom of all the secondary cut holes 17 on the boundary line of the second rhombus 13 and the breaking and decompression explosive sections 173 at the openings of all the secondary cut holes 17 on the boundary line of the third rhombus 14 form a cross-detonation in space and a delayed detonation in time. First, it loosens the upper rock around the secondary cut holes 17 on the boundary line of the third rhombus 14. Then, the explosions of the high-pressure throwing explosive sections 171 at the bottom of all the secondary cut holes 17 on the boundary line of the second rhombus 13 create an upward throwing effect on the rock around the secondary cut holes 17 on the boundary line of the third rhombus 14, thereby improving the explosion effect of the cut area 1 and increasing the tunneling efficiency.

[0056] Sixty milliseconds after controlled detonation, the orifice breaking and pressure relief explosive sections 173 of all the secondary slotted holes 17 on the boundary line of the fourth rhombus 15 detonate simultaneously, producing a similar effect to the orifice breaking and pressure relief explosive sections 173 of all the secondary slotted holes 17 on the boundary line of the third rhombus 14.

[0057] After a 5ms delay, the high-pressure throwing explosive sections 171 at the bottom of all the secondary cut holes 17 on the boundary line of the third rhombus 14 are detonated simultaneously, producing an upward throwing effect on the rocks around the secondary cut holes 17 on the boundary line of the fourth rhombus 15, improving the explosion effect of the cut area 1 and increasing the tunneling efficiency.

[0058] 90ms after controlled detonation, the high-pressure thrown explosive sections 171 at the bottom of all secondary slotting holes 17 on the boundary line of the fourth rhombus 15 are detonated simultaneously. The cumulative blasting effect of the high-pressure thrown explosive sections 171 at the bottom of all secondary slotting holes 17 on the boundary line of the third rhombus 14 is added to further blast, loosen and shatter the rock, so as to achieve the target advance effect of the tunnel.

[0059] In some embodiments, the detonation is performed in the following sequence: initial kinetic energy explosive section 111 of the main slot hole 11 - projectile explosive section 161 of the auxiliary throwing hole 16 - orifice breaking and pressure relief explosive section 173 of the secondary slot hole 17 on the boundary line of the second rhombus 13 - bottom high-pressure throwing explosive section 171 of the secondary slot hole 17 on the boundary line of the second rhombus 13 - orifice breaking and pressure relief explosive section 173 of the secondary slot hole 17 on the boundary line of the third rhombus 14 - bottom high-pressure throwing explosive section 171 of the secondary slot hole 17 on the boundary line of the third rhombus 14 - orifice breaking and pressure relief explosive section 173 of the secondary slot hole 17 on the boundary line of the fourth rhombus 15 - bottom high-pressure throwing explosive section 171 of the secondary slot hole 17 on the boundary line of the fourth rhombus 15. This allows the explosive stress wave generated by the initial detonation to reflect off the wall of the auxiliary throwing hole 16 and form tensile stress in the rock within the rhombus configuration space.

[0060] Traditional shallow-hole blasting suffers from technical bottlenecks such as small cycle advances and frequent operations. Especially in deep mines with high ground stress and complex geological conditions, deep-hole blasting presents challenges in energy transfer and rock fracturing. In some embodiments, a full-face blasting design using a uniform 5m deep hole can be adopted. The lengths of the main cut hole 11 and the secondary cut hole 17 are unified at 5m, breaking through the traditional hole depth of 2.0~3.5m. By using a rhomboid straight-eye cut hole layout, the rhomboid spatial configuration of the cut holes utilizes the geometric interference principle of stress wave propagation. This allows the explosive stress wave generated in the main cut hole 11 to undergo phase reversal at the wall of the auxiliary blast hole 16, resulting in a constructive superposition of the explosive stress wave in the rhomboid spatial configuration. Compared with conventional arrangements, the rhomboid layout can increase the energy concentration of explosive stress waves by 15%~20%, which is particularly suitable for hard rock tunnel excavation in 5m full-face deep holes.

[0061] In some embodiments, the length of the ejection aid hole 16 is greater than the length of the main blast hole 11. The ejection explosive section 161 in the ejection aid hole 16 is set in the extended section at the bottom of the blast hole, which can eject the rock fragments generated after the explosion in the main blast hole 11 upwards, thereby improving the rock breaking and ejection effect in the deep hole blasting zone 1. The extended section at the bottom of the blast hole refers to the blast hole length section in which the length of the ejection aid hole 16 exceeds the length of the main blast hole 11 in the direction away from the hole opening. The effect of setting explosives in the ejection aid hole 16 above the extended section at the bottom of the blast hole on rock breaking and ejection is not obvious. In order to save the use of explosives, the ejection explosive section 161 in the ejection aid hole 16 is only set in the extended section at the bottom of the blast hole.

[0062] In some embodiments, the length of the throwing aid hole 16 is 0.2 to 0.4 m longer than the length of the main slotting hole 11. This provides effective compensation space for subsequent blasting without excessively increasing the drilling workload. The charge at the bottom of the throwing aid hole 16 generates an upward throwing effect, improving the rock breaking and throwing effect in the deep hole slotting area 1.

[0063] In some embodiments, the length of the assisted throwing hole 16 is preferably 0.3m longer than the length of the main scoop hole 11, and two rolls of emulsion explosive are placed at the bottom of the hole to enhance the throwing effect on the rock fragments generated above the assisted throwing hole 16.

[0064] In some embodiments, the main plugging mud section 112 of the main cut hole 11 is 1m long, with the remainder being the length of the explosive charge. The 1m length of the main plugging mud section 112 can effectively seal the hole opening and prevent the explosive gas from escaping prematurely. The remaining hole length is filled with explosive charge to ensure sufficient blasting energy to form the initial cavity, providing ample free face and compensation space for subsequent secondary cut holes, auxiliary holes, and peripheral holes. The length of the main plugging mud section 112 can also be determined based on the hole diameter of the main cut hole 11 and the minimum resistance line theory.

[0065] In some embodiments, the length of the high-pressure thrown explosive section 171 at the bottom of the secondary slot 17 is: L1 = (0.35~0.4) × Hc; the length of the isolation mud section 172 of the secondary slot 17 is: L 泥 =(0.1~0.12)×Hc; Length of the orifice crushing and depressurizing explosive section 173 of the secondary slotting hole 17: L2=(0.25~0.3)×Hc; Length of the secondary plugging mud section 174 of the secondary slotting hole 17: L 堵 =Hc-L1-L 泥 -L2; that is, except for the high-pressure thrown explosive section 171 at the bottom of the hole, the isolation mud section 172, and the secondary plugging mud section 174, all other sections in the secondary slotting hole 17 are secondary plugging mud sections 174; where Hc is the length of the secondary slotting hole 17; the secondary plugging mud section 174 uses mud to plug the borehole; through the above-mentioned segmented charging parameters, the charging ratio of the borehole breaking and decompression explosive section 173 is set to 0.25~0.3, and it explodes before the high-pressure thrown explosive section 171 at the bottom of the same secondary slotting hole 17, pre-breaking the rock mass at the borehole opening of the secondary slotting hole 17, removing the constraint load on the deep rock, removing the borehole opening load, and providing throwing space for the bottom explosive. The space formed provides a spatial basis for the throwing effect generated by the explosion of the high-pressure thrown explosive section 171 at the bottom of the hole, reducing the clamping effect of the rock at the bottom of the borehole; the isolation mud section 172 ensures that the high-pressure thrown explosive section 171 at the bottom of the hole and the borehole are separated. The interval separation of the pressure-relieving explosive section 173 at the borehole opening ensures that the detonation control is strictly implemented according to the specified time interval, avoiding the uncontrollable detonation sequence caused by the explosion of one section triggering a chain reaction of explosions in another section. The charge ratio of the high-pressure throwing explosive section 171 at the bottom of the borehole is set to 0.35~0.4, which can sink the energy center of the explosive explosion to the bottom of the borehole where the ground stress clamping effect is strong, ensuring that the rock at the bottom of the borehole is fully broken and thrown towards the rock fracture space generated by the explosive explosion of the pressure-relieving explosive section 173 at the borehole opening. The sealing length of the secondary plugging mud section 174 extends the working time of the gas wedge effect of the explosive gas on the rock fissures, ensuring that the energy does not leak out and forcibly penetrates into the deep rock mass and rock fissures, effectively sealing the borehole opening to prevent energy dissipation. The coordinated length ratio of each explosive section ensures that the blasting energy is evenly distributed along the depth of the borehole, improving the energy utilization rate of the explosive and the rock breaking effect.

[0066] The lengths of the high-pressure throwing explosive section 171 at the bottom of the hole, the isolation stemming section 172, the hole-opening fragmentation and pressure-relief explosive section 173, and the secondary stemming section 174 in the secondary cut hole 17 are set based on the critical energy matching model derived from the Mohr-Coulomb criterion and the Livingston blasting crater theory in rock mechanics. First, according to the Mohr-Coulomb criterion, the shear strength of deep rocks increases non-linearly with the confining pressure. If the charge length ratio of the high-pressure throwing explosive section 171 at the bottom of the hole is lower than the derived critical value of 0.35, the detonation pressure of the high-pressure throwing explosive section 171 at the bottom of the hole will not be able to overcome the high confining pressure clamping effect at a depth of 5 m, resulting in plastic yield failure of the rock and "root retention". The upper limit of 0.4 prevents excessive comminution of the rock at the bottom of the hole, causing energy waste, and reserves sufficient filling space for the upper isolation stemming section 172, the hole-opening fragmentation and pressure-relief explosive section 173, and the secondary stemming section 174. Second, according to the Livingston blasting crater theory, the length of the hole-opening fragmentation and pressure-relief explosive section 173 needs to meet the optimal burial depth ratio to form an effective fragmentation crater. If the ratio exceeds 0.3, the excess energy generated during the explosion of the hole-opening fragmentation and pressure-relief explosive section 173 will break through the minimum resistance line and be converted into air shock waves and flying rock hazards. If it is lower than 0.25, the upper rock load of the secondary cut hole 17 cannot be removed. Third, according to the theory of the safety distance of explosive sympathetic detonation, if a physical isolation zone of 0.1 - 0.12 is not set, the shock wave generated by the explosion of the high-pressure throwing explosive section 171 at the bottom of the hole will directly induce the sympathetic detonation of the hole-opening fragmentation and pressure-relief explosive section 173, resulting in the failure of the millisecond initiation timing sequence. A precise match between the explosive energy utilization rate and the rock resistance in space is constructed, thus solving the problem that conventional homogeneous charging cannot take into account both deep rock breaking and shallow safety of the blast hole and improving the explosive energy utilization rate. The secondary stemming section 174 is used to adaptively fill the remaining length in the secondary cut hole 17, effectively plugging the hole opening on the premise of meeting the minimum safety resistance line (such as 0.9 m), prolonging the time of the gas wedging effect of the explosion-generated gas on the rock cracks, and improving the explosive energy utilization rate.

[0067] The harder the rock, the longer the charge section is required to overcome the surrounding rock resistance. The softer the rock, excessive charging is likely to cause overbreak, and the charge length needs to be reduced, decreasing / increasing from the harder rock as the reference (i.e., coefficient 1.0) towards both ends. In some embodiments, according to the basic quality index BQ of the rock mass, the lengths of each section of explosive are adjusted. Specifically:

[0068] When BQ > 550, multiply the length of each section of explosive by 1.15 as the final length of each section of explosive;

[0069] When 450 < BQ ≤ 550, multiply the length of each section of explosive by 1.1 as the final length of each section of explosive;

[0070] When 350 < BQ ≤ 450, multiply the length of each section of explosive by 1.0 as the final length of each section of explosive;

[0071] When 250 < BQ ≤ 350, multiply the length of each section of explosive by 0.9 to obtain the final length of each section of explosive;

[0072] When 150 < BQ ≤ 250, multiply the length of each section of explosive by 0.8 to obtain the final length of each section of explosive;

[0073] When BQ ≤ 150, multiply the length of each section of explosive by 0.75 to obtain the final length of each section of explosive.

[0074] For the secondary cut hole 17, when the sum of the final lengths formed by adjusting the charge lengths of each section of explosive according to the above adjustment method and the lengths of the original isolation stemming section 172 and the secondary stemming section 174 is less than 5 m, the length of the secondary stemming section 174 is preferentially increased to fill the secondary cut hole 17; when the sum of the final lengths of each section of explosive charge and the lengths of the original isolation stemming section 172 and the secondary stemming section 174 is greater than 5 m, the length of the isolation stemming section 172 between the bottom high-pressure throwing explosive section 171 and the hole-opening fragmentation pressure-relief explosive section 173 in each secondary cut hole 17 is preferentially shortened, where the length of the isolation stemming section 172 is not less than 0.3 m, to isolate the bottom high-pressure throwing explosive section 171 and the hole-opening fragmentation pressure-relief explosive section 173, preventing the explosion of one section of explosive from triggering the explosion of the other section, ensuring that the explosives in each secondary cut hole 17 are detonated in accordance with the specified initiation sequence and specified time interval, and the length of the secondary stemming section 174 is not less than 0.9 m to ensure the blocking of the explosion-generated blast gases, forming pressure gases in the secondary cut hole 17, thereby producing a fragmentation effect on the rock; when the lengths of both the isolation stemming section 172 and the secondary stemming section 174 reach the minimum lengths and the sum of the lengths of the bottom high-pressure throwing explosive section 171, the hole-opening fragmentation pressure-relief explosive section 173, the isolation stemming section 172, and the secondary stemming section 174 is still greater than 5 m, the length of the blast hole is increased so that all sections of explosive, the isolation stemming section 172, and the secondary stemming section 174 can be placed entirely in the secondary cut hole 17.

[0075] By introducing the basic mass index of rock mass and deriving a dynamic correction coefficient of 0.75 - 1.15 based on the principles of wave impedance matching and conservation of effective crushing work, the technical problem that the "fixed explosive specific consumption" in the prior art cannot adapt to drastic changes in lithology is solved; taking the charging lengths of each section when 350 < BQ ≤ 450 as the wave impedance matching reference value (coefficient 1.0), the length of the explosive is corrected, so that the above blasting parameters can adapt to changes in different geological conditions; for rock masses with BQ > 550 and 450 < BQ ≤ 550, due to stress wave reflection loss caused by wave impedance mismatch, gain coefficients of 1.15 and 1.1 are respectively introduced to compensate for the transmitted energy, ensuring that the effective crushing work is higher than the dynamic compressive strength of the rock; when BQ > 550, the maximum final length of each section of the explosive is set to the length of each section of the explosive × 1.15, anchoring the critical threshold of the effective crushing work of the bottom section of the explosive in the blast holes such as the initial kinetic energy explosive section 111 of the main cut hole 11 and the bottom high-pressure throwing explosive section 171 in the secondary cut hole 17, preventing excessive explosion energy from damaging the geometric interference structure; in the rock section with BQ > 450, the charge amount is appropriately adjusted to the length of each section of the explosive × 1.1, ensuring that there is sufficient blasting energy to overcome the high strength and integrity of the rock and ensuring the crushing effect. The specific increased charge amount is determined by referring to the corresponding range of the above basic mass index BQ of the rock mass and the specific adjustment method of the length of each section of the explosive; for rock masses with 250 < BQ ≤ 350, the dynamic strength of the rock decreases, and a decay coefficient of 0.9 is introduced to avoid energy dissipation in ineffective plastic deformation and over-crushing, appropriately reducing the charge amount to avoid problems such as over-excavation, surrounding rock loosening, and support difficulties caused by excessive blasting energy; for rock masses with 150 < BQ ≤ 250, to prevent premature escape of the explosion-generated gas along the primary fissures and cause the failure of the gas wedge, a correction coefficient of 0.8 is further introduced to reduce the peak pressure and extend the quasi-static action time; for rock masses with BQ ≤ 150, the minimum final length of each section of the explosive is set to the length of each section of the explosive × 0.75, ensuring that the explosion cooperation of each section of the explosive can generate the minimum throwing kinetic energy threshold required to overcome the friction resistance of the 5m deep hole wall when throwing fragmented rock; thus, a deep closed-loop from "axial orderly distribution" to "total amount adaptive regulation" of blasting energy is realized logically, significantly reducing the clamping effect of deep surrounding rock.

[0076] The aforementioned methods for adjusting the length of each explosive segment make the cut-out blasting design more targeted and flexible. They can effectively control the excavation profile quality while ensuring tunneling efficiency, reduce disturbance and damage to the surrounding rock, lower subsequent support costs, and achieve refined adjustments to blasting parameters under different lithological conditions. The same methods for adjusting the length of each explosive segment are applicable to the main cut-out hole 11. These methods supplement the length confirmation scheme for the initial kinetic energy explosive segment 111 in the main cut-out hole 11. Using these methods, the initial kinetic energy explosive segment 111 in the main cut-out hole 11 is adjusted accordingly. The length of section 111 is adjusted. Calculations show that when the length of the main plugging mud section 112 is less than 0.9m, the length of the main plugging mud section 112 must be at least 0.9m to ensure that the length of the main plugging mud section 112 prioritizes meeting the safety requirement of the "minimum resistance line." According to the theory of explosive gas escape, excessively short mud cannot form an effective high-pressure seal. At the moment of explosion, the high-temperature, high-pressure gas will directly rush out of the borehole, causing a "blast," resulting in a sudden drop in the gas wedge pressure inside the borehole, making it impossible to work on deep rock. Therefore, it is necessary to set the length of the main plugging mud section 111... The length of section 12 is not less than 0.9m as a "mandatory safety limit"; that is, when the theoretically calculated charge length results in a remaining space of less than 0.9m, the charge length of the explosive must be reduced to forcibly retain the length of the main plugging mud section 112 of 0.9m; when BQ≤150, the explosive length is the shortest compared to the explosive length under rock conditions with BQ>150, and the length of the main plugging mud section 112 is the longest. Especially for extremely soft rock, the longer length of the main plugging mud section 112 is beneficial to use less explosive energy for "flexible compression" and prevent the rock at the borehole from being excessively thrown or crushed due to excess energy; the length of each section of explosive mentioned above refers to the original theoretical design length of each section of explosive, and the final length of each section of explosive mentioned above refers to the adjusted theoretical length obtained by theoretical calculation based on the original theoretical design length of each section of explosive, taking into account the basic quality indicators of the rock mass and the principle of wave impedance matching and effective breaking work conservation. In specific construction, the calculated theoretical length is processed according to the description in the above scheme to determine the length of each section of explosive and the length of each section of mud in actual construction.

[0077] The basic quality index BQ of the rock mass is calculated according to the formula in the national standard GB / T50218-2014 "Classification Standard for Engineering Rock Mass": BQ=100+3R C +250K V , where R C K represents the saturated uniaxial compressive strength of rock. V R is the rock mass integrity index; where R is the rock mass integrity index. C and K V The values ​​for rock masses with different rock hardness are based on the data in GB / T50218-2014 "Engineering Rock Mass Classification Standard" mentioned above, and will not be repeated here.

[0078] Rocks can be qualitatively classified into hard rocks and soft rocks based on their hardness. Hard rocks are further divided into hard rocks and relatively hard rocks. Hard rocks generally show no water absorption after being submerged; representative examples include granite, syenite, diorite, diabase, basalt, andesite, gneiss, quartz schist, siliceous slate, quartzite, siliceous cemented conglomerate, quartz sandstone, and siliceous limestone. Relatively hard rocks show slight water absorption after being submerged; representative examples include welded tuff, marble, slate, and white marble. Soft rocks include dolomite, limestone, and calcareous cemented sandstone; soft rocks are divided into relatively soft rocks, soft rocks, and extremely soft rocks; relatively soft rocks, after being soaked in water, can be scratched with a fingernail, and representative rocks include tuff, phyllite, sandy mudstone, marl, argillaceous sandstone, siltstone, and shale; soft rocks, after being soaked in water, can be broken apart by hand, including strongly weathered hard rocks, weakly to strongly weathered relatively hard rocks, weakly weathered relatively soft rocks, and unweathered mudstone; extremely soft rocks, after being soaked in water, can be kneaded into a ball, including all kinds of completely weathered rocks and various semi-formed rocks. Classification of rock weathering degree: Unweathered: Structure and texture unchanged, fresh rock; Slightly weathered: Structure and texture, mineral color basically unchanged, some fissure surfaces show iron and manganese inclusions; Weakly weathered: Structure and texture partially destroyed, mineral color changes more obviously, weathered minerals appear on fissure surfaces or weathered interlayers exist; Strongly weathered: Structure and texture mostly destroyed, mineral color changes significantly, feldspar, mica and other minerals are mostly weathered into secondary minerals; Completely weathered: Structure and texture completely destroyed, most mineral components except quartz are weathered into soil-like state.

[0079] In some embodiments, for hard rock, the final length of each explosive segment in the secondary cut hole 17 can be multiplied by 1.1 to increase the explosive length and overcome the surrounding rock resistance; for soft rock, the final length of each explosive segment in the secondary cut hole 17 can be multiplied by 0.9 to reduce the explosive length and avoid over-excavation; when the sum of the final length of each explosive segment and the length of the original isolation mud segment 172 and the secondary plugging mud segment 174 is less than 5m, the length of the secondary plugging mud segment 174 is preferentially increased to fill the secondary cut hole 17; when the sum of the final length of each explosive segment and the length of the original isolation mud segment 172 is less than 5m, the length of the secondary plugging mud segment 174 is preferentially increased to fill the secondary cut hole 17; when the sum of the final length of each explosive segment and the length of the original isolation mud segment 172 is less than 5m, the length of the secondary plugging mud segment 174 is increased to fill the secondary cut hole 17. When the sum of the lengths of section 172 and the secondary plugging mud section 174 is greater than 5m, the length of the isolation mud section 172 between the two explosive sections in each secondary slotting hole 17 is shortened first, wherein the length of the isolation mud section 172 is not less than 0.3m and the length of the secondary plugging mud section 174 is not less than 0.9m; when the lengths of the isolation mud section 172 and the secondary plugging mud section 174 both reach the minimum length, and the lengths of each explosive section and the isolation mud section 172 and the secondary plugging mud section 174 are still greater than 5m, the length of the borehole is increased so that the explosive and the isolation mud section 172 and the secondary plugging mud section 174 can all be placed into the secondary slotting hole 17.

[0080] In some embodiments, the method of adjusting the length of each explosive segment in the secondary slotting hole 17 for different conditions of hard rock and soft rock is also applicable to the main slotting hole 11. That is, for hard rock, the final length of the initial kinetic energy explosive segment 111 in the main slotting hole 11 is the original design length of the initial kinetic energy explosive segment 111 × 1.1; for soft rock, the final length of the initial kinetic energy explosive segment 111 in the main slotting hole 11 is the original design length of the initial kinetic energy explosive segment 111 × 0.9.

[0081] In some embodiments, the charge density of the initial kinetic energy explosive section 111 of the main slotting hole 11 is 1.15~1.35 kg / m; the detonator 2 of the main slotting hole 11 is disposed at the bottom of the main slotting hole 11; the charge density of the high-pressure throwing explosive section 171 at the bottom of the secondary slotting hole 17 is 1.15~1.35 kg / m, and the charge density of the breaking and depressurizing explosive section 173 at the orifice is 0.95~1.15 kg / m; the charge density of the high-pressure throwing explosive section 171 at the bottom of the secondary slotting hole 17 and the orifice... All sections of the crushing and decompression explosive section 173 are equipped with detonators 2 at their bottom; the detonation delay between the high-pressure throwing explosive section 171 at the bottom of the hole and the crushing and decompression explosive section 173 at the orifice is 40ms; the initial kinetic energy explosive section 111 of the main slotting hole 11 uses high-density charge, which can form high-intensity detonation pressure in the deep area of ​​the deep hole where the clamping force is strong, forcibly tearing the rock at the bottom of the blast hole of the main slotting hole 11. When the integrity of the rock is destroyed and the broken rock leaves its original position, it reduces the clamping force of the surrounding rock and provides initial space for subsequent blasting.

[0082] For the initial kinetic energy explosive section 111 of the main cut hole 11 and the high-pressure thrown explosive section 171 at the bottom of the secondary cut hole 17, this application abandons the conventional density charge mode (usually <1.15kg / m) commonly used in existing deep hole blasting, and instead adopts a high-density charge of 1.15~1.35kg / m. The conventional density charge commonly used in the prior art has a low detonation wave impedance, which cannot overcome the clamping effect of the rapidly increasing ground stress at the bottom of a 5m deep hole, and is very prone to the "rooting" phenomenon caused by insufficient detonation pressure leading to the rock at the bottom of the hole not being broken. However, this application has a high density charge mode for the initial kinetic energy explosive section 111 and the high-pressure thrown explosive section 171 at the bottom of the hole. The charge density of the explosive section 171 adopts an unconventional range of 1.15~1.35 kg / m. This density is derived from the "critical fracturing density" based on the dynamic compressive strength of deep rock under triaxial high confining pressure. The detonation energy density of the explosive at this density is significantly improved, which can form a high-intensity detonation pressure sufficient to offset the confining pressure in the area with the strongest deep confining force, thereby forcibly tearing the rock at the bottom of the borehole. When the integrity of the rock is destroyed and the rock fragments are removed from their original positions, the confining force of the surrounding rock can be significantly reduced and initial space can be provided for subsequent blasting, thus achieving a better deep-hole rock breaking effect than existing conventional density charge technology.

[0083] During the explosion, the orifice-breaking and pressure-relief explosive section 173 of the secondary cut hole 17 removes the load on the rock at the orifice section of the secondary cut hole 17 with appropriate energy, reducing the energy consumption of the high-pressure throwing explosive section 171 at the bottom of the hole. The charge density of the high-pressure throwing explosive section 171 at the bottom of the secondary cut hole 17 is the same as that of the initial kinetic energy explosive section 111 of the main cut hole 11. On the one hand, it shatters the rock at the bottom of the secondary cut hole 17, and on the other hand, it throws the rock fragments generated by the explosion of the orifice-breaking and pressure-relief explosive section 173 out of the position of the secondary cut hole 17, providing more space for subsequent rock blasting. At the same time, by utilizing the spatiotemporal superposition of the explosive stress waves generated by the explosion of the high-pressure throwing explosive section 171 at the bottom of the hole and the orifice-breaking and pressure-relief explosive section 173, the explosive energy in the deep hole is transformed from "disordered impact" to "directional pushing", thereby maximizing rock breaking efficiency while ensuring construction safety.

[0084] See Figure 3 In some embodiments, considering the energy distribution characteristics of a 5m deep hole, the charge density of the reinforced charge section at the bottom of the 5m main cut hole 11 is 1.15~1.35kg / m, and the charge length is 1.5m, which is used to resist the clamping effect at the bottom of the deep hole. Since the bottom of the hole is located deep in the rock mass, the triaxial stress constraint is large and the clamping effect is strong, requiring a high charge density to provide concentrated blasting energy. Combined with the rock fragmentation space formed by the blasting at the top of the hole, it ensures that the rock at the bottom of the hole is fully fragmented and forms an effective initial slot cavity, providing compensation space and free surface for the collapse of the upper rock and the blasting of the rock in adjacent blast holes.

[0085] The charge density of the conventional charge section in the middle is 0.95~1.15 kg / m, and the charge length is 2.5m. It is used to maintain the stable crushing of rocks within the corresponding range of the deep hole. The rock clamping effect in the middle section is relatively weakened. At the same time, the rock crushing space formed by the explosion of adjacent boreholes further reduces the clamping effect of the surrounding rocks. A lower charge density can achieve uniform crushing, while avoiding excessive rock crushing caused by excessive energy concentration, thus improving the effective utilization rate of explosive energy.

[0086] The main plugging mud section 174 at the top of the borehole is 1.0m long. The main plugging mud section 174 at the top of the borehole makes full use of the pressure of the explosive gas generated in the deep hole on the rock wall to create cracks and break rocks. The reasonable length of the main plugging mud section 174 at the borehole can effectively seal the borehole, prolong the action time of the explosive gas in the borehole, allow the high-pressure gas to fully expand and do work, enhance the gas wedge effect and breaking effect on the rock around the borehole wall, prevent the explosion energy from escaping from the borehole too early, control the distance of the flying rocks, and ensure construction safety.

[0087] With the above charging structure, a differential distribution of blasting energy along the hole depth direction is achieved, which matches the stress characteristics and fragmentation requirements of the rocks at various parts of the 5-m deep hole. The high energy of the explosive in the bottom enhanced charging section is concentrated for breakthrough, the distribution of the explosive in the middle conventional charging section is uniform, and the energy steadily fragments. The secondary plugging stemming section 174 at the hole mouth utilizes gas pressure to assist in crack formation. The three sections act synergistically to improve the overall blasting efficiency of the deep hole cut, improve the forming quality of the cut cavity, and create good free face conditions for the subsequent blasting of other cut holes, auxiliary holes, and perimeter holes.

[0088] In some embodiments, the blasting method further includes the step of分级 setting the spacing between the cut holes according to the lithological conditions: setting the spacing based on the integrity of the rock mass (applicable to areas where the development of joints and fractures has a significant impact on blasting). According to the rock mass quality index RQD or the number of joint sets Jn, determine the spacing S between the cut holes:

[0089] Take a larger spacing in areas with relatively intact lithology and appropriately reduce the spacing in joint-developed areas. Specifically:

[0090] When the rock mass quality index RQD≥85% or the number of joint sets Jn≤2 sets, the spacing S between the cut holes is calculated according to the formula: ,

[0091] The calculation is carried out, and the value range is 900~1120mm;

[0092] When 60%≤RQD<85% or 2<Jn≤3 sets, the spacing S between the cut holes is calculated according to the formula: ,

[0093] The calculation is carried out, and the value range is 680~900mm;

[0094] When RQD<60% or Jn>3 sets, the spacing S between the cut holes is calculated according to the formula: ,

[0095] The calculation is carried out, and the value range is 400~680mm; where RQD is the rock mass quality index, RQD refers to the percentage of the cumulative length of intact rock cores with a length≥10cm in the drilled rock cores to the total length of the footage; Jn is the number of joint sets, S is the spacing between the cut holes, d is the diameter of the blast hole, D is the detonation velocity of the explosive, ρ e [[ID=`35]]is the density of the explosive, σ c is the uniaxial compressive strength of the rock.

[0096] The slotting holes here include main slotting holes 11 and secondary slotting holes 17; the slotting hole spacing refers to the distance between two adjacent secondary slotting holes 17 and the distance between the main slotting hole 11 and the adjacent secondary slotting hole 17. By classifying and setting the slotting hole spacing according to lithological characteristic parameters such as rock mass quality index RQD, joint group number Jn, rock firmness coefficient f, and joint spacing, blasting parameters can be optimized for different lithological conditions. This effectively avoids uneven blasting caused by differences in rock mass integrity, achieving sufficient fragmentation in hard and intact rock masses while preventing over-blasting in fractured rock masses. This improves blasting efficiency and fragmentation quality, reduces construction costs, and enhances the safety and controllability of blasting operations.

[0097] RQD and Jn are two core indicators in rock mechanics for evaluating rock mass integrity. Physically, they are negatively correlated, but their influence on blasting effects is consistent. For example, RQD ≥ 85% means a long intact section in the core and low rock fragmentation; Jn ≤ 2 means few joint groups and structural planes in the rock mass, approximating a intact medium. Both "high RQD" and "low Jn" physically represent good rock mass integrity and few defects. In such rock masses, the scattering and attenuation of explosive stress waves are slowest, and the effective range is largest. Therefore, although their mathematical signs are opposite, their physical orientation is the same. Thus, the rock mass quality index RQD or the number of joint groups Jn can be used to determine the spacing between cut holes. When the rock mass quality index RQD≥85% or the number of joint groups Jn≤2 groups, the constant term "200" in the calculation formula of the slot hole spacing S is a physical compensation for the slower attenuation of explosive stress waves and the larger effective work radius in rock masses with RQD≥85% or the number of joint groups Jn≤2 groups. This avoids the explosive energy of the explosive in the slot hole that explodes first affecting the explosive effect of the explosive in the slot hole that explodes later due to the excessively small spacing between the slot holes. At the same time, it improves the utilization of explosive energy and reduces explosive waste. The critical hole spacing coefficient of traditional shallow holes is generally 1.5~2.0, while the 5m deep hole working condition in this application requires the use of a larger stress wave superposition field. After correcting the critical hole spacing coefficient of traditional shallow holes, 2.5 is adopted as the critical hole spacing coefficient of the slot hole. Physically proportional to the detonation pressure of explosives, it represents the intensity of the shock wave generated at the moment of the explosion, i.e., the active attack force attempting to shatter rocks. A larger value indicates greater explosive energy and a larger destructive radius. σ represents the detonation pressure of an explosive, used to quantify the force required to break rocks; c Let d be the uniaxial compressive strength of the rock, representing the essential property of the rock to resist fragmentation; that is, the higher the uniaxial compressive strength of the rock, the more difficult it is to be blasted apart. Here, d is the diameter of the borehole. 2 As a geometric factor in the denominator, it is mainly used to correct for size effects. In deep-hole blasting, as the hole diameter increases, the rock's resistance to fracturing and its clamping effect changes non-linearly. The rock resistance representing the rock-bound pore size factor is used to quantify the resistance of the rock to the energy of the explosive; the ratio of the two A dimensionless matching coefficient representing the relative relationship between the rock-breaking ability of the explosive and the rock resistance. When increases, that is, the explosive energy is greater or the rock is easier to be broken by the explosive, it means that the explosive energy is in an advantageous state relative to the rock resistance, and the radius of the fracture circle generated by the explosion is larger. Therefore, the spacing S between two adjacent cut holes can be set larger, and it can still ensure the fracture connection between two adjacent cut holes. On the contrary, when decreases, that is, the explosive energy is weak or the rock is difficult to be broken by the explosive, it means that the rock is difficult to be blasted open, and the range of fracture expansion generated by rock fragmentation is limited. Therefore, it is necessary to reduce the cut hole spacing S and use the superimposed intensive explosion energy to forcibly break the rock; according to the aspherical geometric divergence of the explosion stress wave and the non-linear attenuation law of the explosion energy with distance propagation, set the stress wave attenuation exponent of to 0.25; in addition, by limiting the value range of the cut hole spacing S to 900 - 1120 mm, it is ensured that sufficient stress can be generated between two adjacent cut holes to fragment the rock, and at the same time, it is prevented that the explosion energy is overly concentrated, resulting in ineffective energy consumption for pulverizing the rock; at the same time, to avoid the influence of the explosion energy between adjacent cut holes and increase the cross-sectional area of the roadway formed when the cut hole spacing S is increased, considering from the economic aspect, since the cross-sectional area of the roadway in roadway tunneling will not expand infinitely, when the explosive energy in the cut hole is too large, by limiting the value range of the cut hole spacing S, the borehole diameter d of the cut hole can be inversely limited, and by reducing the borehole diameter of the cut hole, the charge amount of the explosive in the borehole can be reduced, and to a certain extent, the release of the explosive energy can be reduced, and the influence between the explosives in two adjacent cut holes during explosion can be reduced.

[0098] When 60% ≤ RQD < 85% or 2 < Jn ≤ 3 groups, in the calculation formula of the cut hole spacing S, based on ensuring moderate explosive energy consumption hedging and standard radius regression, the additional 200 mm is removed from the formula, making the stress wave generated by the explosion of the explosive in the cut hole be affected by joint scattering in the rock mass of 60% ≤ RQD < 85% or 2 < Jn ≤ 3 groups and the physical reality that the conduction performance of the explosion stress wave returns to normal; at the same time, the value range of the cut hole spacing S is limited to 680 - 900 mm, moderately shrinking the hole spacing relative to RQD ≥ 85% or the number of joint groups Jn ≤ 2 groups, actively hedging the wave energy loss caused by the structural plane, and ensuring that the stress wave generated by the explosion of the explosive in the cut hole realizes the connection between adjacent cut holes before decaying to the critical value, so as to still ensure the ability to break the rock required for deep hole cutting in moderately intact rock masses or under complex joint conditions.

[0099] When RQD < 60% or Jn > 3 groups, in the calculation formula of the cut hole spacing S, based on the compensation of the scattering of explosive energy and the guidance of the weak planes of rock fractures, introducing the constant term "-150" can compensate for the reflection and scattering losses of the explosive energy on a large number of natural joint surfaces, ensuring that under the condition that the effective action radius of the stress wave in the discontinuous medium with a large number of joint groups is significantly reduced, the explosion effect can still be guaranteed; the value range of the cut hole spacing S is limited to 400 - 680 mm, further shrinking the hole spacing compared with 60% ≤ RQD < 85% or 2 < Jn ≤ 3 groups, ensuring that the energy of the superimposed stress of the high-density stress field generated by multiple simultaneously exploding cut holes can break the rock mass before being completely absorbed by the rock weak planes; the gaps between the fragmented rocks generated in the cut holes that explode first provide moving space for the movement generated during the fragmentation of the rocks in subsequent explosions, which is beneficial to the fragmentation of the rocks in the subsequent explosion area, and the explosion-generated explosion gases clear the fragmented rocks out of the cut holes, providing a larger moving space for the movement of the rocks caused by the crushing force of the subsequent explosion on the rocks.

[0100] It is also possible to set the spacing based on the rock hardness. In some embodiments, the blasting method further includes the step of determining the cut hole spacing S according to the rock solidity coefficient f or the joint spacing:

[0101] In the intact rock mass area where the rock solidity coefficient f ≥ 12 or the joint spacing > 1.5 m, the value range of the cut hole spacing S is 400 - 600 mm;

[0102] In the moderately intact rock mass area where the rock solidity coefficient 8 ≤ f < 12 or the joint spacing is 0.5 - 1.5 m, the value range of the cut hole spacing S is 600 - 800 mm;

[0103] In the fractured rock mass area where the rock solidity coefficient f < 8 or the joint spacing < 0.5 m, the value range of the cut hole spacing S is 800 - 1120 mm;

[0104] Among them, f is the rock solidity coefficient, and S is the spacing between the cut holes.

[0105] The rock solidity coefficient characterizes the hardness of the rock. Although the rock being solid does not directly indicate the integrity of the rock structure, for the unmined rock area, a high rock solidity means that the rock is difficult to be damaged in its original state, so it indirectly indicates the integrity of the rock structure in the blasting area. The joint spacing directly characterizes the integrity of the rock structure. In the dynamics of blasting rock breaking, both can show the characteristics of "high blast resistance" of the rock; the larger the rock solidity coefficient, the greater the "cohesive strength" generated by the mineral bonding of the rock, and a larger joint spacing means that the rock mass lacks the "original weak planes" available for blasting and crushing the rock. Both lead to the difficulty of breaking the rock during blasting. Therefore, the cut hole spacing S can be set based on the rock solidity coefficient and the joint spacing.

[0106] In rock mass regions where the rock strength coefficient f ≥ 12 or joint spacing > 1.5m, the rock mass exhibits extremely high wave impedance and lacks primary weak surfaces. The radial crack propagation radius induced by single-hole blasting is typically only 15 to 18 times the hole diameter. Therefore, setting the upper limit of the slot hole spacing S to 600mm aims to ensure that the stress wave peak superposition area of ​​adjacent blast holes covers the midpoint of the line connecting adjacent blast holes, thereby overcoming the high cohesive strength of the rock and achieving rock fragmentation between two adjacent slot holes and forced penetration between two adjacent blast holes. The lower limit of 400mm can avoid deep hole drilling deviation. The risk of physical perforation is mitigated by preventing the dynamic plastic locking effect of rock caused by the small spacing between two adjacent blast holes. This means that the explosion compresses and densifies the rock rather than stretches and fractures it, thus ensuring the formation of an effective ultra-high pressure superimposed stress field within a range of 400-600mm. The value of the slot hole spacing S is 400-600mm, ensuring that the explosive energy of the explosive in the closer slot holes in the intact rock mass is "concentrated" to generate an ultra-high pressure superimposed stress field sufficient to break the solid bedrock, thereby breaking the rock and reducing the surrounding rock confinement effect in the rock area of ​​subsequent explosions.

[0107] In rock mass areas where the rock strength coefficient is 8≤f<12 or the joint spacing is 0.5~1.5m, the value of the slot hole spacing S is 600~800mm. By utilizing the existing gaps between the rocks, the rock fragmentation effect in the first blasting hole releases some of the surrounding rock stress, reducing the clamping effect of the deep hole surrounding rock on the subsequent blasting area. While maintaining the above-mentioned spacing range between each slot hole without weakening the explosive energy in the slot hole, the explosive energy can be fully utilized, and the clamping effect of the surrounding rock can be reduced.

[0108] In rock mass regions where the rock firmness coefficient f < 8 or the joint spacing < 0.5m, the value of the slot hole spacing S ranges from 800 to 1120mm. By utilizing the characteristic that the fractured rock mass is prone to fluctuation under the action of explosive stress wave and explosive gas, the explosive energy is used to drive the fractured rock blocks to move and be thrown. This ensures that even if the slot hole spacing is increased, the rock can still be broken and thrown, providing sufficient space for rock movement in subsequent explosions.

[0109] In some embodiments, when determining the spacing of the cut holes, the RQD most directly reflects the overall quality of the rock mass, and the number of joint groups Jn affects the propagation path of the explosive stress wave. Therefore, the RQD is first used as the priority judgment standard, and the basic spacing is determined according to the above calculation formula for the spacing of the cut holes S. Then, it is finely adjusted according to the number of joint groups Jn (±50~100mm). Finally, it is corrected according to the rock firmness coefficient f, and the spacing of the cut holes is multiplied by a coefficient of 0.9~1.1 to obtain the final determined spacing of the cut holes.

[0110] In some embodiments, the diameters of the main slotting hole 11, the blasting aid hole 16, and the secondary slotting hole 17 are determined based on rock parameters; specifically,

[0111] The diameter of the main slot hole 11 is calculated according to the formula:

[0112] ,

[0113] Perform calculations;

[0114] The diameter of the 16mm diameter hole for the assisted polishing is calculated using the following formula: ,

[0115] Perform calculations;

[0116] The diameter of the inner layer slotted hole 17 is calculated according to the formula: ,

[0117] Perform calculations;

[0118] The diameter of the outer layer slotted hole 17 is calculated according to the formula: ,

[0119] Calculations are performed; where D1, D2, D3, and D4 are the diameters of the main slotting hole 11, the auxiliary polishing hole 16, the inner layer slotting hole 17, and the outer layer slotting hole 17, respectively. It represents the uniaxial compressive strength of the rock. 1 is the rock density; RQD is the rock mass quality index, ranging from 0 to 100%; f is the rock firmness coefficient; Jn is the joint group number; K5 is the deep hole correction coefficient for the main cut hole, ranging from 1.15 to 1.25; K6 is the compensation correction coefficient for the blasting hole, ranging from 0.9 to 1.1; K7 is the breakage correction coefficient for the inner layer cut hole, ranging from 0.95 to 1.05; K8 is the transmission correction coefficient for the outer layer cut hole, ranging from 0.9 to 1.0; the inner layer cut hole 17 is the secondary cut hole 17 on the boundary of the second rhombus 13 and the third rhombus 14; the outer layer cut hole 17 is the secondary cut hole 17 on the boundary of the fourth rhombus 15. By establishing a borehole diameter calculation system based on multiple parameters such as rock uniaxial compressive strength, rock density, rock mass quality index, rock firmness coefficient, and joint group number, and by setting corresponding correction coefficients for boreholes with different functions such as main slotting holes, auxiliary blasting holes, inner layer slotting holes, and outer layer slotting holes, it is possible to achieve precise matching between borehole diameter and rock properties, ensuring that boreholes at each level play their best role in the blasting process.

[0120] The diameter of the blast hole 11 is based on 32mm, and the uniaxial compressive strength of the rock is introduced. Power function of density A semi-empirical, semi-theoretical formula was constructed based on dimensional analysis and a rock blasting breaking energy index model; among which, as well as It is a dimensionless normalization process in physics, aiming to establish a universal benchmark for rock blastability, namely, the threshold compressive strength of typical hard rock of 100 MPa and the average density of crustal rocks of 2.6 g / cm³. 3 As a reference calibration value, specific lithological parameters are converted into relative blast resistance coefficients relative to standard rocks, thereby eliminating the interference of different physical dimensions on the calculation weights. The power exponent in the formula reflects the nonlinear mechanical characteristics of the energy required for rock destruction, characterizing the sensitivity of the borehole diameter to the geometrical increase in rock resistance. That is, the harder the rock, the more the required borehole diameter needs to increase not only proportionally but also exponentially to provide sufficient work volume. The adjustment coefficient K5 is an engineering correction factor used to map the theoretically calculated continuous values ​​to the discrete values ​​of the industrial standard drill bit diameter sequence and to compensate for the systematic errors caused by on-site geological defects and drilling equipment wear, ensuring the feasibility of the calculation results in engineering practice. This quantifies the physical resistance of the bedrock to the expansion of the borehole wall of the main cut hole 11. The greater the uniaxial compressive strength of the rock, the harder the rock, and the greater the rock density, the larger the diameter of the main cut hole 11 needs to be to accommodate higher-energy explosives, thereby generating sufficient initial detonation pressure to break the original rock structure under high ground stress.

[0121] The borehole diameter of the blasting aid hole 16 is set to be 1.6 to 2 times that of the main cut hole 11, and is positively correlated with the rock mass quality index (RQD). The constant 1.6 in the formula defines the critical diameter ratio threshold for the formation of the effective free surface in the blasting aid hole 16. Its physical meaning lies in the fact that, according to deep-hole blasting theory, only when the diameter of the empty hole reaches at least 1.6 times that of the charging hole can the tangential tensile stress concentration factor formed around it effectively overcome the closing effect of the deep ground stress in the borehole. The rock mass quality index (RQD) was dimensionlessly normalized and converted into a rock integrity coefficient ranging from 0 to 1. A coefficient of 0.4 represents the maximum fragmentation space compensation factor, physically indicating that as rock mass integrity increases, the fragmentation coefficient after rock fracturing increases. Therefore, it is necessary to linearly increase the volume of the blasting hole to provide additional expansion compensation space; that is, for every 10% increase in RQD, an additional 0.04 times the hole diameter compensation is required. (1.6 and...) The sum physically constitutes the theoretical optimal borehole enlargement ratio, which is formed by the superposition of the basic free surface requirement and the dynamic fragmentation requirement. The multiplication operation of K6 with the above sum is physically a mapping of engineering gradation. Through overall scaling, the calculated continuous theoretical multiple is forcibly mapped and matched to the discrete drill bit diameter sequence of the industry standard, ensuring that the final selected blasting hole diameter meets the theoretical expansion requirement while being in the optimal matching range of the existing drill torque and slag removal capacity. Utilizing the free surface provided by the blasting hole 16, when the RQD is higher, by enlarging the borehole diameter, on the one hand, the wave impedance mismatch caused by the large-diameter blasting hole generates a strong reflected tensile wave, causing the rock on the hole wall to flake and shatter; on the other hand, the enlarged physical cavity volume directly provides sufficient geometric displacement and fragmentation compensation space for the broken rock, preventing the broken rock from being "crushed" again by the surrounding rock due to insufficient movement space after fracturing.

[0122] In the calculation of the diameter of the inner layer slotted hole 17, a logarithmic function is introduced. This allows the aperture of the inner layer slotting hole 17 to be nonlinearly adjusted with the rock firmness coefficient f. The inner layer slotting hole 17 is located between the main slotting hole 11 and the outer layer slotting hole 17. When the explosive stress wave of the inner layer slotting hole 17 and its inner side diffuses outward, a smooth transition is made according to the brittle characteristics of the rock corresponding to the rock firmness coefficient f. This avoids the explosive stress wave from being ineffectively scattered at the rock layer interface due to the sudden change in the aperture of the inner layer slotting hole 17 during propagation, thereby ensuring the regularity of the groove formed by the explosion in the area near the blast hole.

[0123] In the calculation of the diameter of the outer layer cut hole 17, based on the calculation of the diameter of the inner layer cut hole 17, a reduction correction is made by Jn. The outer layer cut hole 17 faces a large number of joint groups, which makes the loss of explosive energy more serious. The more joint groups there are, the greater the negative shrinkage adjustment of the borehole diameter needs to be, that is, to further reduce the borehole diameter, thereby reducing the single-hole charge and detonation peak pressure, and preventing high-pressure gas from escaping too early along the joint surface, which would cause the gas wedge effect to fail. The borehole diameter is adjusted according to Jn to balance the breaking power of the explosive energy and the crack guidance of the rock. The end of the explosive energy is used to accurately cut off the last connection between the cut area 1 and the surrounding rock, and eliminate the rock clamping effect.

[0124] In the diameter calculation formulas of the main cut hole 11, the auxiliary blasting hole 16, the inner layer cut hole 17, and the outer layer cut hole 17, a hierarchical calculation model is constructed with the main cut hole 11 as the fixed energy source, the auxiliary blasting hole 16 as the expansion source, and the inner and outer layer cut holes 17 as the adjustment source. The borehole diameter is coupled with rock mechanical parameters, including uniaxial compressive strength, rock density, rock mass quality index, and rock firmness coefficient, as well as joint group characteristics, and fine-tuned using correction coefficients. Among them, K5 counteracts the high ground stress resistance at the depth of the borehole through energy gain, K6 is used to dynamically adjust the reflected wave intensity and spatial compensation rate of the auxiliary blasting hole 16, K7 ensures the smooth transmission of stress waves between the inner layer cut holes 17 to optimize the uniformity of rock fragmentation, and K8 precisely controls the convenience of blasting cutting through energy convergence. The entire parameter model calculation system precisely controls the intensity of the explosive stress wave through the geometric interface of the differentiated borehole layout, ensuring the dynamic balance between the explosive fragmentation energy and the compensation space under deep hole conditions.

[0125] As the length of the main cut hole 11 increases, the resistance of the surrounding rock at the depth of the main cut hole 11 gradually increases. In order to increase the ability to discharge broken rock, in some embodiments, the extension length of the auxiliary discharge hole 16 compared to the main cut hole 11 is determined according to the rock mass quality index RQD: when RQD≥75%: extension length = 0.08×H z +150mm; When 25%≤RQD<75%: Extension length = 0.06×H z +100mm; When RQD < 25%: Extension length = 0.04 × H z +50mm; where H zThe length of the main cut hole 11 is given. Based on the rock mass quality index RQD classification, the length of the auxiliary throwing hole 16 is set to be longer than that of the main cut hole 11. The length of the proportional coefficient × Hz represents the length of the explosive section 161, and the constant term represents the ultra-deep void space at the bottom of the hole. That is, the spatial distribution of the auxiliary throwing hole 16 from the hole opening to the bottom is: no explosive space - explosive section 161 - no explosive space at the bottom of the hole represented by the constant term. When RQD ≥ 75%, the rock mass is generally relatively intact, with a correspondingly large stress clamping effect, high tensile strength, and tight rock block interlocking. The unit volume of explosive consumption required to eject deep rock, i.e., the specific energy consumption, is the same as that required for fractured rock. Since the length of the main cut hole 11 is more than twice that of the main body, an upper limit coefficient of 0.08 is set so that the proportional coefficient of the lengthening of the auxiliary throwing hole 16 increases linearly with the length of the main cut hole 11. That is, for every 1m increase in the depth of the main cut hole 11, the auxiliary throwing hole 16 needs to add an extra 8cm of charge length to provide sufficient throwing kinetic energy to throw the rock fragments generated by the explosion of the explosive in the main cut hole 11 and the auxiliary throwing hole 16 out of the blast hole. In rocks with RQD≥75%, the bottom of the blast hole is the strongest "dead corner" of the surrounding rock clamping force. A 150mm ultra-deep space is reserved as an "energy-concentrating cavity" to concentrate high-power explosives to crush the rock root at the bottom of the hole, prevent bottom residue, and ensure The minimum physical space required for the explosive in the extended section of the assisted-explosive hole 16 ensures the ability of the explosive section 161 to expel rock fragments during detonation. When RQD < 25%, the rock mass is relatively fragmented, with a loose structure and low self-weight stress, making it extremely easy to eject. Using a high coefficient would lead to uncontrolled flyrock, so the proportional coefficient is lowered to 0.04. When RQD < 25%, the rock at the bottom of the hole generally has developed fissures or is relatively soft, without strong "dead zones." A constant of 50mm is sufficient to cover the space required for detonation, preventing over-drilling due to the rock mass already having multiple weak surfaces. When 25% ≤ RQD < 75%... When the proportional coefficient is set to the midpoint between 0.08 and 0.04, i.e., 0.06, and the constant term is set to the midpoint between 150mm and 50mm, i.e., 100mm; in the above formula, the specific energy consumption coefficient and the ultra-deep margin constant term are dynamically adjusted to match the throwing resistance characteristics of rocks with different lithologies, so as to achieve a balance between the throwing power generated by the explosion of the projectile explosive section 161 in the blasting hole 16 and the construction energy consumption. Under the premise of ensuring sufficient cutting space, excessive drilling is avoided, which not only improves the uniformity and controllability of blasting effect, but also optimizes construction efficiency and economy, and provides a scientific parameter design basis for precise blasting under complex geological conditions.

[0126] In some embodiments, the blasting method further includes: the first rhombus 12 having at least one projection-aiding hole 16 at each vertex; the second rhombus 13, the third rhombus 14, and the fourth rhombus 15 having at least one secondary slotting hole 17 at each vertex; the vertices of the first rhombus 12 coinciding with the midpoints of the sides of the second rhombus 13; the vertices of the second rhombus 13 coinciding with the midpoints of the sides of the third rhombus 14; the vertices of the third rhombus 14 coinciding with the midpoints of the sides of the fourth rhombus 15; each vertex of the first rhombus 12 having a projection-aiding hole 16, such that the projection-aiding holes 16 are uniformly and symmetrically distributed along the diagonal of the first rhombus 12; secondary slotting holes 17 being provided at the vertices of the second rhombus 13, the third rhombus 14, and the fourth rhombus 15, so that during detonation, the explosive stress waves from the secondary slotting holes 17 of the third rhombus 14 and the fourth rhombus 15 are uniformly superimposed; the second rhombus 13... The high-pressure throwing explosive section 171 at the bottom of the secondary slotting holes 17 on the third rhombus 14 and the fourth rhombus 15, and the breaking and decompression explosive section 173 at the orifice, explode to form a continuous stress wave transmission chain, improving the utilization rate of explosive energy. The vertex of the first rhombus 12 coincides with the midpoint of the side of the second rhombus 13; the vertex of the second rhombus 13 coincides with the midpoint of the side of the third rhombus 14; and the vertex of the third rhombus 14 coincides with the midpoint of the side of the fourth rhombus 15. This causes the various throwing holes 16 and secondary slotting holes 17 to form regularly distributed explosion points along the first rhombus 12, the second rhombus 13, the third rhombus 14, and the fourth rhombus 15. During detonation, this enhances the superposition effect of the explosive stress waves between the various throwing holes 16 and secondary slotting holes 17. The stress waves between the blast hole at the midpoint of each rhombus's side and the blast holes symmetrically distributed on both sides of the midpoint are compressed and superimposed, improving the explosive blasting effect.

[0127] In some embodiments, the vertex of the first rhombus 12 is on the side of the second rhombus 13; the vertex of the second rhombus 13 is on the side of the third rhombus 14; and the vertex of the third rhombus 14 is on the side of the fourth rhombus 15. Depending on the rock distribution, the blasting aid hole 16 at the vertex of the first rhombus 12 is positioned on the side of the second rhombus 13. If the rock distribution is uneven, the blasting aid hole 16 at the vertex of the first rhombus 12 is also asymmetrically distributed on the side of the second rhombus 13. The vertices of the second rhombus 13 and the third rhombus 14 are also distributed according to the distribution pattern of the blasting aid hole 16 at the vertex of the first rhombus 12 on the side of the second rhombus 13. This facilitates targeted blasting of harder rocks and optimizes the superposition effect of explosive stress waves between each blast hole.

[0128] In some embodiments, the first rhombus 12, the second rhombus 13, the third rhombus 14, and the fourth rhombus 15 are all squares, which facilitates the positioning of blast holes and improves the uniformity of the distribution of the blasting aid holes 16 and the secondary slotting holes 17. When the explosives in the blast holes on each rhombus boundary are detonated, the explosive stress waves generated by the explosives in each blast hole are uniformly superimposed in the rock of the rhombus configuration space, which enhances the stress wave superposition effect and improves the explosion effect.

[0129] Example 2

[0130] See Figure 4 This application provides a method for full-section 5m deep-hole slotting blasting in rock tunnel excavation. The tunnel cross-section in this embodiment is a three-centered arch shape, with a width of 4600mm, a height of 3500mm, a sidewall height of 3100mm, and an arch height of 400mm. The surrounding rock is mainly biotite granulite and biotite amphibolite schist, with a rock hardness coefficient f<15, belonging to the semi-hard to hard rock category, and a saturated compressive strength of 29.9~146.96MPa, providing typical geological conditions for the application of 5m deep-hole blasting technology.

[0131] First, boreholes were drilled to sample the rock to be excavated, and laboratory tests were conducted to determine the rock's physical properties and basic mechanical parameters. Two standard rock sample sizes were used: 50×100mm and 50×25mm. Physical properties such as density, porosity, and water content were tested. Key indicators for rock mechanical parameters included uniaxial compressive strength, tensile strength, elastic modulus, Poisson's ratio, and angle of internal friction. This provides a theoretical basis and data support for the design of blasting parameters for a 5m deep hole.

[0132] During the blasting parameter design phase, considering the characteristics of 5m deep-hole full-section blasting design, the blasting parameters for cut holes, caving holes, peripheral holes 4, and bottom holes 5 were determined, including the number of blast holes, their arrangement, and the selection of explosives. An optimized 5m deep-hole full-section layout scheme was adopted. Cut zone 1 is located in the central area of ​​the tunnel end face. Cut zone 1 is equipped with 13 cut holes and 4 auxiliary throwing holes 17. The 13 cut holes include 1 main cut hole 11 and 12 secondary cut holes 17. The 4 auxiliary throwing holes 17 and 12 secondary cut holes 17 are arranged in a diamond pattern. The caving zone 3 is arranged around cut zone 1. The caving zone 3 is equipped with 26 caving holes, and 12 peripheral holes 4 and 11 bottom holes 5 are arranged in the peripheral forming area outside the caving zone 3. The spacing between each blast hole is controlled within the range of 400~700mm. The length of the secondary blast hole is 200~500mm longer than that of the main blast hole. The spacing between the blasting zone 1 and the collapse zone 3 and the bottom hole 5 is 0.8~1.2m. The spacing between the collapse holes adjacent to the perimeter holes 4 in the collapse zone 3 and the perimeter holes 4 is 0.8~1.2m. After the blasting zone 1 is detonated, the blast holes of the collapse holes, perimeter holes 4 and bottom hole 5 in the collapse zone 3 are detonated in sequence, gradually expanding outward along the radial direction with the main blast hole 11 as the center.

[0133] One main cut hole 11 is located in the center of the cut area 1, and adopts a 5m deep hole design; see Figure 2 The main slot hole 11 adopts an axial gradient charging structure, with the bottom 1.5m being a reinforced charging section; the middle 2.5m being a conventional charging section; and the top being a secondary plugging mud section 174, which uses mud and water bags to seal the hole, with a sealing length of not less than 1m, to ensure effective confinement of the explosive gas; the detonator 2 is a digital electronic detonator, with a delay accuracy controlled within ±1ms.

[0134] The 12 secondary cut holes 17 are the same length as the main cut hole 11 and are arranged at the vertices of the second rhombus 13, the third rhombus 14, and the fourth rhombus 15, respectively. The secondary cut holes 17 employ segmented charging technology, with the secondary cut holes 17 being charged in two sections, upper and lower. A separating section of stemming material 172 separates the high-pressure throwing explosive section 171 at the bottom of the hole from the crushing and depressurizing explosive section 173 at the opening, with a spacing length of not less than 0.5m. The top of the secondary cut holes 17, i.e., the part near the borehole opening, is further reinforced with a secondary plugging section of stemming material 174. The process involves blocking the flow of explosives. The lower section refers to the portion near the bottom of the hole, and the upper section refers to the portion near the opening. During detonation, the upper explosive detonates first, followed by the lower explosive. There is a 40ms delay between the detonation of the upper and lower explosives to ensure favorable blasting conditions for the lower explosive. The first detonation of the upper explosive provides a free surface and compensation space for the lower explosive, effectively overcoming the clamping effect of the surrounding rock under 5m deep hole conditions. Segmented blasting achieves a reasonable distribution of energy within the deep hole, avoiding ineffective fragmentation caused by excessive energy concentration.

[0135] The auxiliary ejection hole 16 is located at the apex of the first rhombus 12, centered on the main slotting hole 11. The depth of the auxiliary ejection hole 16 is 0.3m longer than the lengths of the main slotting hole 11 and the secondary slotting hole 17. The 0.3m extended section at the bottom of the hole holds φ32mm emulsion explosive, providing compensation space for the explosion of the surrounding slotting holes. At the same time, the explosive section 161 at the bottom generates an upward throwing effect during the explosion, improving the rock fragmentation and throwing effect in the 5m deep slotting area 1. The diameter of the auxiliary ejection hole 16 is 90mm, and the diameter of the charging hole is 38mm. The diameter of the auxiliary ejection hole 16 and the charging hole form a 2.37-fold diameter ratio. Utilizing the phase reversal effect of stress waves on the wall of the auxiliary ejection hole 16, the incident compression wave is converted into a reflected tension wave. The spacing between adjacent auxiliary ejection holes 16 is controlled within the range of 300~560mm, realizing the synergistic superposition effect of the reflected waves from multiple auxiliary ejection holes 16, reducing the clamping effect of the surrounding rock.

[0136] During the drilling phase, rock drilling equipment such as fully hydraulic rock drilling rigs and umbrella drill frames were used, equipped with hollow four-corner drill rods with a 5.5m drill bit. A laser-guided drilling rig and a fixed guide frame at the borehole opening were used to ensure drilling accuracy control. The slotting holes, auxiliary holes, and peripheral holes were drilled sequentially according to the design. After drilling the blast holes, the depth of each hole was precisely measured to determine if it met the design requirements. Simultaneously, each hole was thoroughly cleaned to ensure that no impurities were present that would affect the charging effect.

[0137] During the charge structure implementation phase, the boreholes were divided into three categories: slotted holes, auxiliary holes, and peripheral holes, totaling 13 slotted holes. A specialized segmented charging structure within the holes and a bottom-reinforced charging technology for the 16th ejection aid hole were employed. No. 2 rock emulsion explosive was used, with a diameter of 32mm, a detonation velocity of 4300m / s, and a density of 1200kg / m³. The slotted holes adopted an axial gradient charging structure: a 1.5m reinforced charge section at the bottom with a charge density of 1.25kg / m³, a 2.5m conventional charge section in the middle with a charge density of 1.05kg / m³, and a 1.0m secondary plugging mud section at the top. The 16th ejection aid hole was 0.3m deeper than the other holes, reaching 5.3m in length, with a diameter of 90mm. Two rolls of emulsion explosive, each weighing 0.3kg, were placed at the bottom of the hole. The slotted area 1 adopts the in-hole segmented charging technology, with the upper and lower sections charged separately. The sections are separated by a 172-segment isolation mud section. The upper section detonates first and the lower section detonates later, with a 40ms delay between sections. This reduces blasting vibration and safety risks, and facilitates the improvement of the stability and controllability of the overall blasting effect.

[0138] In the detonation control stage, millisecond-series delayed electric detonators are used, and the detonation sequence adopts a three-stage design: In the first stage, after detonation, the projectile explosive sections 161 in the main slotting hole 11 and the auxiliary throwing hole 16 are detonated sequentially. The explosive stress wave generated by the detonation of the main slotting hole 11 is reflected by the wall of the auxiliary throwing hole 16 and forms tensile stress in the rock within the rhomboid construction space of the first rhomboid 12. The stress waves between the simultaneously detonated slotting holes are superimposed, improving the rock fragmentation effect between the simultaneously detonated slotting holes. The second rhomboid 13... The projectile explosive section 161 in the apex of the throwing hole 16 detonates simultaneously, causing an upward throwing effect on the broken rock in the main cutting hole 11, expanding the initial space; in the second section, the explosives in the secondary cutting holes 17 at the apex of the second rhombus 13 and the third rhombus 14 detonate with a 50ms delay, achieving stress wave superposition; in the third section, the explosives in the secondary cutting hole 17 at the apex of the fourth rhombus 15 detonate with a 110ms delay, achieving quasi-static throwing; through precise timing control, it is easy to improve the delayed breaking quality and throwing effect.

[0139] It should be noted that, in this document, relational terms such as "first" and "second" are used only to distinguish one entity or operation from another, and do not necessarily require or imply any such actual relationship or order between these entities or operations. Furthermore, the terms "comprising," "including," or any other variations thereof are intended to cover non-exclusive inclusion, such that a process, method, article, or apparatus that comprises a list of elements includes not only those elements but also other elements not expressly listed, or elements inherent to such a process, method, article, or apparatus. Without further limitations, an element defined by the phrase "comprising one..." does not exclude the presence of other identical elements in the process, method, article, or apparatus that includes said element. The terms “center,” “longitudinal,” “lateral,” “length,” “width,” “thickness,” “upper,” “lower,” “front,” “rear,” “left,” “right,” “vertical,” “horizontal,” “top,” “bottom,” “inner,” “outer,” “clockwise,” “counterclockwise,” “axial,” “radial,” and “circumferential” indicate the orientation or positional relationship based on the orientation or positional relationship shown in the accompanying drawings. They are used only for the convenience of describing the embodiments of this application and simplifying the description, and are not intended to indicate or imply that the device or element referred to must have a specific orientation, or be constructed and operated in a specific orientation. Therefore, they should not be construed as limitations on the embodiments of this application.

[0140] The above description is merely a specific implementation of the embodiments of this application, but the protection scope of the embodiments of this application is not limited thereto. Any variations or substitutions that can be easily conceived by those skilled in the art within the technical scope disclosed in the embodiments of this application should be included within the protection scope of the embodiments of this application. Therefore, the protection scope of the embodiments of this application should be determined by the protection scope of the claims.

Claims

1. A method for full-section deep-hole blasting in rock shafts, characterized in that, Including the steps: Set main cut holes in the cut area; Uniformly distribute multiple auxiliary throwing holes on the boundary line of the first rhombus centered on the main cut hole; Outside the first rhombus, from the main cut hole outwards, there are successively a second rhombus, a third rhombus and a fourth rhombus centered on the main cut hole; and multiple secondary cut holes are uniformly distributed on the boundary lines of the second rhombus, the third rhombus and the fourth rhombus, and the secondary cut holes have the same length as the main cut hole; Set explosives in the main cut holes, secondary cut holes and auxiliary throwing holes; where, in the main cut hole, from the bottom of the hole to the top of the hole, there are successively: an initial kinetic energy explosive section and a main stemming clay section; in the secondary cut hole, from the bottom of the hole to the top of the hole, there are successively: a bottom-hole high-pressure throwing explosive section, an isolation clay section, a hole-mouth fragmentation pressure-relief explosive section and a secondary stemming clay section; set a throwing explosive section at the bottom of the auxiliary throwing hole; Initiate simultaneously according to the following order: the initial kinetic energy explosive section of the main cut hole and the hole-mouth fragmentation pressure-relief explosive sections of all the secondary cut holes on the boundary line of the second rhombus - the throwing explosive sections of all the auxiliary throwing holes on the boundary line of the first rhombus - the bottom-hole high-pressure throwing explosive sections of all the secondary cut holes on the boundary line of the second rhombus and the hole-mouth fragmentation pressure-relief explosive sections of all the secondary cut holes on the boundary line of the third rhombus - the bottom-hole high-pressure throwing explosive sections of all the secondary cut holes on the boundary line of the third rhombus and the hole-mouth fragmentation pressure-relief explosive sections of all the secondary cut holes on the boundary line of the fourth rhombus - the bottom-hole high-pressure throwing explosive sections of all the secondary cut holes on the boundary line of the fourth rhombus.

2. The blasting method according to claim 1, characterized in that, The lengths of the auxiliary throwing holes are all greater than the length of the main cut hole, and the throwing explosive sections in the auxiliary throwing holes are set in the bottom extended section of the blast hole.

3. The blasting method according to claim 1, wherein The length of the bottom-hole high-pressure throwing explosive section of the secondary cut hole: L1 = (0.35 - 0.4) × Hc; The length of the isolation mud section of the secondary cut hole: L 泥 =(0.1~0.12)×Hc; The length of the hole-mouth fragmentation pressure-relief explosive section of the secondary cut hole: L2 = (0.25 - 0.3) × Hc; The length of the secondary plugging mud section of the secondary slotting hole: L 堵 =Hc-L1-L 泥 -L2; Where, Hc is the length of the secondary cut hole.

4. The blasting method according to claim 3, characterized in that, According to the basic quality index BQ of the rock mass, adjust the lengths of each section of explosives, specifically: When BQ > 550, multiply the length of each section of explosives by 1.15 as the final length of each section of explosives; When 450 < BQ ≤ 550, multiply the length of each section of explosives by 1.1 as the final length of each section of explosives; When 350 < BQ ≤ 450, multiply the length of each section of explosives by 1.0 as the final length of each section of explosives; When 250 < BQ ≤ 350, multiply the length of each section of explosives by 0.9 as the final length of each section of explosives; When 150 < BQ ≤ 250, multiply the length of each section of explosives by 0.8 as the final length of each section of explosives; When BQ ≤ 150, multiply the length of each section of explosives by 0.75 as the final length of each section of explosives.

5. The blasting method according to claim 1, characterized in that, The charge density of the initial kinetic energy explosive section of the main cut hole is 1.15 - 1.35 kg / m; the detonator of the main cut hole is arranged at the bottom of the main cut hole; the charge density of the high-pressure throwing explosive section at the bottom of the secondary cut hole is 1.15 - 1.35 kg / m, and the charge density of the hole-opening fragmentation pressure-relief explosive section is 0.95 - 1.15 kg / m; detonators are arranged at the bottoms of both the high-pressure throwing explosive section and the hole-opening fragmentation pressure-relief explosive section at the bottom of the secondary cut hole; the detonation delay between the high-pressure throwing explosive section at the bottom of the hole and the hole-opening fragmentation pressure-relief explosive section is 40 ms.

6. The blasting method according to claim 1, characterized in that, It also includes the steps: Classify and set the spacing between the cut holes according to the lithological conditions: When the rock mass quality index RQD≥85% or the number of joint sets Jn≤2 sets, the spacing S between the cut holes is calculated according to the formula: ， and the value range is 900 - 1120 mm; When 60%≤RQD<85% or 2<Jn≤3 sets, the spacing S between the cut holes is calculated according to the formula: ， and the value range is 680 - 900 mm; When RQD<60% or Jn>3 sets, the spacing S between the cut holes is calculated according to the formula: ， and the value range is 400 - 680 mm; Where RQD is the rock mass quality index, Jn is the number of joint groups, and S is the spacing between the cut holes. The diameter of the borehole. For the detonation velocity of the explosive, For the density of the explosive, It represents the uniaxial compressive strength of the rock.

7. The blasting method according to claim 1, characterized in that, It also includes the steps: Determine the spacing S between the cut holes according to the rock hardness coefficient f or the joint spacing: In the intact rock mass area where the rock hardness coefficient f≥12 or the joint spacing>1.5 m, the value range of the spacing S between the cut holes is 400 - 600 mm; In the moderately intact rock mass area where the rock hardness coefficient 8≤f<12 or the joint spacing is 0.5 - 1.5 m, the value range of the spacing S between the cut holes is 600 - 800 mm; In the fractured rock mass area where the rock hardness coefficient f<8 or the joint spacing<0.5 m, the value range of the spacing S between the cut holes is 800 - 1120 mm; where, f is the rock hardness coefficient and S is the spacing between the cut holes.

8. The blasting method according to claim 1, characterized in that, The diameters of the main cut hole, the assisting throwing hole and the secondary cut hole are determined according to the rock parameters; specifically, The diameter of the main cut hole is calculated according to the formula: ， for calculation; The diameter of the assisting throwing hole is calculated according to the formula: ， for calculation; The diameter of the inner-layer secondary cut hole is calculated according to the formula: ， for calculation; The diameter of the outer-layer secondary cut hole is calculated according to the formula: ， Perform the calculations; where D1, D2, D3, and D4 are the diameters of the main slotting hole, the auxiliary blasting hole, the inner layer slotting hole, and the outer layer slotting hole, respectively. It represents the uniaxial compressive strength of the rock. ρ is the rock density; RQD is the rock mass quality index, ranging from 0 to 100%; f is the rock firmness coefficient; Jn is the joint group number; K5 is the main slotting hole depth correction coefficient, ranging from 1.15 to 1.25; K6 is the auxiliary blasting hole compensation correction coefficient, ranging from 0.9 to 1.1; K7 is the inner layer slotting hole breakage correction coefficient, ranging from 0.95 to 1.05; K8 is the outer layer slotting hole transmission correction coefficient, ranging from 0.9 to 1.0; the inner layer slotting hole is the secondary slotting hole on the boundary of the second and third rhombuses; the outer layer slotting hole is the secondary slotting hole on the boundary of the fourth rhombus.

9. The blasting method according to claim 1, characterized in that, It also includes: At least one assisting throwing hole is arranged at each vertex of the first rhombus, and at least one secondary cut hole is arranged at each vertex of the second rhombus, the third rhombus and the fourth rhombus; The vertex of the first rhombus coincides with the midpoint of the side of the second rhombus; the vertex of the second rhombus coincides with the midpoint of the side of the third rhombus; the vertex of the third rhombus coincides with the midpoint of the side of the fourth rhombus.

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