Optimization method of double plugging and graded vibration reduction blasting for compound wedge cut of extremely hard rock tunnel

Abstract

The application discloses a kind of extremely hard rock tunnel compound wedge cut double plugging grading vibration reduction blasting optimization method, belong to the field of tunnel engineering drilling and blasting method construction technology.The application adopts "central hole+first-order cut hole+second-order cut hole" three-level compound wedge cut structure, designs double plugging segmented charging system, sets up grading accurate delay initiation network, establishes low equipment-dependent visual parameter closed-loop optimization method, realizes the improvement of explosive energy utilization rate and blasting vibration accurate control, reaches the unity of efficient rock breaking and micro-vibration control, especially suitable for uniaxial compressive strength ≥200MPa, Protodyakonov's firmness coefficient f≥20 high-integrity extremely hard rock tunnel scene.

Description

Technical Field

[0001] This invention belongs to the field of tunnel engineering drill and blast construction technology, specifically involving an optimized method for graded vibration reduction blasting in extremely hard rock tunnels using a double-sealing, compound wedge-shaped slotting technique. Background Technology

[0002] Extremely hard rock formations are characterized by high uniaxial compressive strength, dense rock mass, poor joint and fissure development, and poor drillability and blastability. Drill-and-blast method remains the mainstream construction method for tunnels in this type of extremely hard rock. In drill-and-blast construction of extremely hard rock tunnels, the effectiveness of the cut-out directly determines the tunnel's cycle advance and blasting efficiency, while the control of blasting vibration directly relates to the stability of the surrounding rock and the safety of surrounding buildings and structures. These are the two core control challenges in blasting construction of extremely hard rock tunnels.

[0003] Currently, a series of mature technical solutions have been developed for slotting optimization, charge structure design, and micro-vibration control in tunnel blasting. For ordinary hard and medium-hard rock formations with a Protodyakonov coefficient f=10~18, core parameters such as slotting hole angle, hole spacing, and depth ratio have been optimized, resulting in standardized design methods adapted to medium-hard rock formations. Regarding charge structure and micro-vibration control, existing technologies mostly employ segmented charge structures such as water-interval charges and air-interval charges, combined with high-precision micro-delay technology using electronic detonators, to achieve graded release of blasting energy and peak offset cancellation of vibration waves.

[0004] However, when the above-mentioned existing technical solutions are directly applied to the blasting construction of extremely hard rock tunnels with uniaxial compressive strength ≥200MPa and f≥20, there are significant compatibility defects, specifically in the following aspects:

[0005] First, the existing cut structure parameters are not compatible with the characteristics of extremely hard rock formations, making it difficult to guarantee the cutting effect. The depth ratio and borehole network parameters of existing compound wedge cuts are optimized for medium-hard and ordinary hard rock formations. However, extremely hard rock formations have extremely high integrity, and the resistance to rock fragmentation and fracture propagation is more than twice that of medium-hard rocks. The existing cut parameters cannot form a sufficient and stable initial free surface, resulting in a strong clamping effect at the bottom of the cut hole, which easily leads to cutting failure and insufficient cycle advance.

[0006] Secondly, existing charging and sealing structures are ill-suited to the high-energy concentrated release characteristics under extremely hard rock blasting conditions, resulting in low explosive energy utilization and poor vibration reduction and dust suppression effects. In extremely hard rock formations, the rock integrity is high, the bottom clamping effect is strong, and the explosive energy required for rock fracturing is highly concentrated. Existing spaced charging structures are prone to energy dissipation, insufficient free surface formation, and low deep rock breaking efficiency. Furthermore, existing technologies often employ a single stemming plugging structure, which has insufficient energy-locking capacity at the borehole opening and cannot withstand the high internal pressure generated by extremely hard rock blasting. This easily leads to borehole overflow, premature leakage of explosive gases, and dust escape, which not only reduces explosive energy utilization but also increases peak blasting vibration and deteriorates the construction environment.

[0007] Third, the existing micro-delay design does not match the formation law of the free surface of extremely hard rock, which easily leads to excessive vibration superposition. When the existing delay parameters are directly applied to extremely hard rock scenarios, the problem of subsequent blast holes being detonated before the previous blast hole has formed a stable free surface is easily encountered. This not only fails to achieve the graded utilization of rock-breaking energy, but also causes the blasting vibration waves to superimpose in the same direction, with the peak vibration velocity far exceeding the safety threshold, seriously threatening the stability of the tunnel surrounding rock and the safety of surrounding buildings and structures.

[0008] Fourth, existing technical solutions have high implementation barriers and are difficult to adapt to the on-site construction conditions of mountain tunnels. Existing blasting parameter optimization schemes rely heavily on specialized equipment and techniques such as numerical simulation software, professional blasting vibration testing instruments, and laboratory rock mechanics tests. However, extremely hard rock tunnels are mostly located in remote mountainous areas, where on-site construction teams generally lack specialized equipment and technical personnel, making it impossible to complete complex parameter calculations and testing optimizations. On-site construction relies heavily on worker experience, resulting in poor stability of blasting effects and a high risk of over- or under-excavation, excessive vibration velocity, and insufficient cycle advance, making it difficult to achieve standardized and regulated construction.

[0009] In summary, the blasting process in extremely hard rock is a continuous process involving the formation of free surfaces, the transfer of explosive energy, and the gradual fracturing of the rock mass. A single technical measure cannot simultaneously meet the demands of space formation, energy utilization, and fracturing control. Therefore, it is necessary to design the cut structure, energy confinement method, and detonation sequence collaboratively from the perspective of the entire blasting process to achieve overall optimization of the blasting effect under extremely hard rock conditions. Currently, there is no integrated blasting construction method for high-integrity, high-strength extremely hard rock tunnel scenarios that balances cut rock-breaking efficiency and blasting vibration control, while also possessing low landing threshold and strong on-site adaptability. Therefore, to address the above issues, developing a compound wedge-shaped cut double-sealing staged vibration-reducing blasting method suitable for extremely hard rock tunnels, and solving the shortcomings of existing technologies such as poor cut effect, low energy utilization, difficulty in vibration control, and poor landing adaptability, has become an urgent technical problem to be solved in this field. Summary of the Invention

[0010] This application aims to address at least one of the technical problems existing in the prior art. To this end, the present invention provides an optimized method for graded vibration reduction blasting in extremely hard rock tunnels using a double-seal, double-wedge-cut structure. This method optimizes the double-wedge-cut structure, designs a double-seal segmented charging system, sets up a graded, precise, delayed detonation network, and establishes a low-equipment-dependency, visualized parameter closed-loop optimization method. This achieves improved explosive energy utilization and precise control of blasting vibration, unifying efficient rock breaking and micro-vibration control. It is particularly suitable for high-integrity extremely hard rock tunnel scenarios with a uniaxial compressive strength ≥200MPa and a Protodyakonov robustness coefficient f ≥20.

[0011] This invention is designed around the continuous process of "space formation - energy action - fragmentation and propagation" in the blasting of extremely hard rock. It establishes initial free space through a compound wedge-shaped cut, improves the efficiency of explosive energy action through a double-sealing structure, and controls the sequence of explosive energy release through graded delays, thus ensuring the interconnection of each technical step and achieving synergistic optimization of the blasting process under extremely hard rock conditions. The specific technical solution adopted is as follows:

[0012] An optimized method for staged vibration reduction blasting in the double-seal, compound wedge-shaped cut-out stage in extremely hard rock tunnels includes the following steps:

[0013] Step 1: Determine the parameters of the extremely hard rock mass

[0014] On-site samples were taken from the rock mass to be blasted to measure its basic physical and mechanical parameters, including rock density. Rock longitudinal wave velocity Rock static Poisson's ratio Uniaxial compressive strength of rock Rock elastic modulus Rock Protodextrin coefficient ƒ and rock tensile strength ;

[0015] Step 2: Calculate the key parameters of the slotting hole

[0016] A three-stage compound wedge-shaped arrangement of "central borehole + first-stage cut hole + second-stage cut hole" is adopted. There are 3 central boreholes arranged vertically along the tunnel's central axis, 8 first-stage cut holes, and 10 second-stage cut holes. The core parameters are determined according to the following formula:

[0017] (1) Vertical depth of the second-order slotted hole: = H + d, where H is the tunnel design cycle advance, and d is the over-depth value, d=0.2~0.4m;

[0018] (2) Vertical depth of first-order slotted hole: H1 = η·H2, where η = H1 / H2 = 0.55~0.65;

[0019] (3) Total charge per hole: Q = q·S·L·ɑ, where q is the empirical value of charge per unit in extremely hard rock, S is the bearing area of ​​a single hole, L is the depth of the hole, and ɑ is the correction coefficient for extremely hard rock.

[0020] (4) Segmented charge ratio: λ=Q1 / Q2=2 / 3, where Q1 is the charge amount at the orifice opening and Q2 is the charge amount at the orifice bottom, and Q1=Q×40% and Q2=Q×60%;

[0021] Step 3: Calculate the fill length of the double-sealing system's stemming material.

[0022] Construct a dual-sealing system consisting of an "in-hole isolation sealing layer + orifice composite sealing layer," with the in-hole isolation sealing layer having a length of [missing information]. =(8~12)· ,in The borehole diameter; the total length of the composite sealing layer at the borehole opening. The following conditions must be met: (1) Not less than 1 / 3 of the borehole depth; (2) Not less than 0.3m; (3) Not less than the minimum resistance line W of the borehole.

[0023] The orifice composite sealing layer consists of, from the inside out, a high-efficiency water gun mud layer, a high-density gun mud layer, and a polyurethane foam layer.

[0024] Step 4: Calculation of Time Delay Parameters for Grading Extremely Hard Rock

[0025] Design a three-stage detonation sequence:

[0026] Level 1: In-hole graded delay, the charging section at the detonation orifice in the center hole and the second-stage cut-out hole has a delay Δ. Then, the explosive charge section at the bottom of the detonation hole;

[0027] Level 2: Inter-row graded delay, the detonation sequence is: the first-order slotted hole integral charge section and the center hole opening section → the second-order slotted hole opening section → the center hole and the bottom section of the second-order slotted hole.

[0028] Level 3: Staggered detonation in the same row, inter-hole detonation of adjacent blast holes in the same ring of auxiliary holes, and inter-hole detonation of adjacent blast holes in the same row of peripheral blast holes.

[0029] Step 5: Verification of Single-Stage Explosive Charge

[0030] The maximum permissible amount of explosive charge per stage is calculated based on the Sadovsky empirical formula. The total amount of explosive charge for all stages detonated simultaneously, calculated in Step 2, is compared with the maximum permissible amount of explosive charge per stage. If the requirements are not met, adjustments are made.

[0031] Step 6: Determining the Explosion Effect and Optimizing Parameter Cycle

[0032] The blasting effect is determined by visual indicators, and the blasting parameters are iteratively optimized until two consecutive blasting cycles meet the design requirements.

[0033] Specifically, the parameters of the slotted hole mesh in Step 2 are as follows:

[0034] The diameter of the center hole is the same as the diameter of the blast hole, and the depth is the same as that of the second-order slotting hole;

[0035] The inclination angle of the first-stage slotting holes is 65°~70°, the hole spacing is 0.4~0.5m, and the row spacing is 0.5~0.6m;

[0036] The inclination angle of the second-order slotting holes is 75°~80°, the hole spacing is 0.5~0.6m, and the row spacing is 0.6~0.7m;

[0037] The minimum distance between the slotted hole and the auxiliary hole is 0.8~1.0m.

[0038] Specifically, in Step 3:

[0039] The length of the isolation and sealing layer inside the borehole is 8 to 10 times the borehole diameter in ordinary extremely hard rock scenarios, and 10 to 12 times the borehole diameter in extremely hard rock scenarios.

[0040] The length of the mud layer in the high-efficiency water cannon is 1 / 10 to 1 / 5 of the borehole depth. Its formula is water and 0.5% to 1% surfactant, wherein the surfactant is detergent or laundry powder.

[0041] The thickness of the polyurethane foam layer is 25~35cm, and the curing time is ≥15min;

[0042] High-density blasting mud layer is compacted and filled in layers to the total sealing length. .

[0043] Specifically, in Step 4: the first-stage hole-level delay Δ The second-level inter-row delay Δ is 55-60ms. The delay time is 50~70ms, and the second-level staggered peak delay is 3~4ms. All delay parameters are rounded up to an integer multiple of the minimum programmable delay of the electronic detonator.

[0044] Specifically, in Step 5, the formula for predicting the peak vibration velocity of a mass point using the general Sadovsky empirical formula from the "Safety Regulations for Blasting" GB 6722-2014 can be modified to obtain:

[0045] Q max = ;

[0046] In the formula: Q max[v] is the maximum permissible single-stage detonation charge; [v] is the safe permissible peak particle vibration velocity; R is the distance from the detonation center; K and α are site coefficients, with K = 50–150 under extremely hard rock conditions. =1.3~1.8;

[0047] When the total charge of explosives detonated simultaneously during any level of delay is less than or equal to the maximum permissible charge of explosives in a single stage, the detonation parameter is deemed to meet the blasting vibration safety requirements. When the total charge of explosives detonated simultaneously during any level of delay is greater than the maximum permissible charge of explosives in a single stage, the method of correction is to reduce the number of blast holes detonated simultaneously in the same stage, adjust the order of graded delays, or reduce the charge of explosives in a single hole until the blasting vibration safety requirements are met.

[0048] Specifically, in Step 6, the core indicators for judging the blasting effect include: ① Hole utilization rate: measure the length of the remaining blast hole and calculate the hole utilization rate η = (H - remaining hole length) / H × 100%; ② Large block rate: randomly select three 1m × 1m × 1m cubic areas in the middle of the blast pile and count the proportion of rock blocks larger than 30cm; ③ Surrounding rock stability and vibration: observe whether there are falling blocks or cracks in the surrounding rock of the tunnel and determine the stability of the surrounding rock; ④ Sealing effect: observe whether there is any blasting at the borehole opening or leakage of blasting gas.

[0049] The parameter optimization rules for cases that do not meet the standards are as follows: ① If the borehole utilization rate is <93%: increase the proportion of the charge section at the bottom of the borehole, or extend the segmented delay Δ in the borehole. ② If the large block ratio is >8%, shorten the time delay Δ between the slotted hole and the auxiliary hole row. ③ If rockfall occurs and there is significant tremor in the surrounding area: increase the inter-row delay Δ ④ If blasting or excessive dust occurs at the orifice: increase the total length of the composite sealing layer at the orifice. Alternatively, the polyurethane foam sealing layer can be thickened.

[0050] Specifically, the termination condition for the iterative cycle in Step 6 is: two consecutive blasting cycles must meet the design requirements of borehole utilization rate ≥93%, large block rate ≤8%, surrounding rock stability, and no impact blasting.

[0051] Compared with existing technologies, this invention is based on a synergistic design of rock-breaking mechanism in extremely hard rock bursting. The three elements create a synergistic effect on the following two levels:

[0052] Parameter correlation level: The charge structure and spatial parameters determined in Step 2 provide the basic conditions for the sealing length design in Step 3 and the delay scheme in Step 4. The energy retention effect in Step 3 and the timing control in Step 4 together promote the gradual expansion of the blasting space, completing the integration of energy locking and energy use. The three are interlinked.

[0053] At the functional synergy level: Step 2 solves the problem of "free surface formation", Step 3 solves the problem of "energy retention", and Step 4 solves the problem of "timing optimization" - the three steps correspond to the three breakpoints in the vicious cycle of ultra-hard rock blasting, and together break the cycle of "strong clamping effect → insufficient free surface → low energy utilization rate → excessive vibration".

[0054] The specific technical effects of the above-mentioned collaborative relationship are manifested in the following ways:

[0055] (1) In view of the characteristics of the high clamping effect of extremely hard rock, the present invention designs a compound wedge-shaped slotting structure with a central hole, optimizes the vertical depth ratio of the first-order and second-order slotting holes to 0.55 to 0.65, optimizes the bottom charge ratio of the second-order slotting holes and the central hole to 60%, and, in conjunction with the step-by-step detonation sequence, effectively induces stress concentration and gradually weakens the clamping effect at the bottom of the hole. The utilization rate of the blast hole in the field is steadily increased from 75% to 80% of the traditional method to more than 93%, and the cyclic advance is increased by 15% to 20%.

[0056] (2) This invention constructs a dual-sealing system of "in-hole isolation sealing + orifice composite sealing", clarifies the quantitative parameters and construction technology of each sealing layer, effectively prevents the two-stage charge detonation in the hole, prolongs the action time of the explosive gas in the hole, increases the energy utilization rate of the explosive by more than 20%, and effectively reduces the risk of blasting at the orifice; the combination of high-efficiency water-based gun mud with added surfactant and polyurethane foam for sealing significantly reduces dust concentration and CO concentration, and greatly shortens the tunnel ventilation time.

[0057] (3) The present invention designs a three-level detonation sequence with in-hole graded delay (55-60ms), inter-row graded delay (50-70ms), and same-row staggered peak delay (3-4ms), which realizes the staggered cancellation of vibration waves and the orderly release of explosion energy. Under the premise of ensuring the advance, the peak value of blasting vibration is reduced by 25%-35% compared with the traditional method, which meets the safety requirements of surrounding sensitive buildings (structures).

[0058] (4) This invention is based entirely on classical theoretical formulas and parameters verified by field tests. It does not rely heavily on numerical simulation software and professional vibration testing equipment. It establishes a visual parameter closed-loop optimization method. On-site construction personnel only need to observe and measure to complete parameter adjustment. It has a low threshold for implementation, strong engineering adaptability, and can realize the rapid determination and standardized operation of blasting construction parameters for extremely hard rock tunnels. Attached Figure Description

[0059] To more clearly illustrate the technical solutions in the embodiments of the present invention, the accompanying drawings used in the description of the embodiments will be briefly introduced below. Obviously, the accompanying drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0060] Figure 1 This is a flowchart of a method according to an embodiment of the present invention;

[0061] Figure 2 This is a schematic diagram of the planar arrangement of the compound wedge-shaped slotted blast holes according to an embodiment of the present invention;

[0062] Figure 3 This is a schematic diagram of the longitudinal cross-sectional arrangement of the compound wedge-shaped slotted blast holes according to an embodiment of the present invention;

[0063] Figure 4 This is a schematic diagram of the layered structure of the composite sealing layer inside the hole according to an embodiment of the present invention;

[0064] Figure 5 This is a flowchart of the graded delayed detonation sequence according to an embodiment of the present invention. Detailed Implementation

[0065] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0066] In the description of this invention, it should be understood that 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 this invention 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 this invention.

[0067] Furthermore, the terms "first" and "second" are used for descriptive purposes only and should not be construed as indicating or implying relative importance or implicitly specifying the number of technical features indicated. Thus, a feature defined as "first" or "second" may explicitly or implicitly include one or more of that feature. In the description of this invention, "a plurality of" means two or more, unless otherwise explicitly specified.

[0068] See Figure 1 An optimized method for staged vibration reduction blasting in the complex wedge-shaped cut-out double-sealing stage in extremely hard rock tunnels includes the following steps:

[0069] Step 1: Determine the parameters of the extremely hard rock mass

[0070] On-site samples were taken from the rock mass to be blasted to measure its basic physical and mechanical parameters, including rock density. Rock longitudinal wave velocity Rock static Poisson's ratio Uniaxial compressive strength of rock Rock elastic modulus Rock Protodextrin coefficient ƒ and rock tensile strength .

[0071] Step 2: Calculate the key parameters of the slotting hole

[0072] A three-stage compound wedge arrangement of "center hole + first-stage slotting hole + second-stage slotting hole" is adopted, such as... Figure 2 As shown.

[0073] There are three central holes, arranged vertically along the tunnel's central axis. The hole depth is the same as that of the second-stage cut holes. A segmented charging structure is adopted for the hole opening section and the hole bottom section to form the initial free surface and enhance the deep cut effect.

[0074] There are a total of 8 first-order slotting holes, 4 on each side, symmetrically arranged around the center hole, forming an inner layer of compound wedge-shaped structure. Their vertical depth is less than that of the second-order slotting holes, and they are used to form the initial slotting cavity.

[0075] There are a total of 10 second-stage slotting holes, 5 on each side, symmetrically arranged around the first-stage slotting holes, forming an outward-expanding compound wedge structure. The depth of the holes is the same as that of the central hole, which is used to achieve deep slotting and rock breaking at the bottom of the hole.

[0076] Determine the core parameters using the following formula:

[0077] 1. Structural parameter calculation

[0078] First, determine the vertical depth of the second-order cut hole. The vertical depth of the second-order cut hole is determined based on the tunnel design cycle advance and the over-depth value:

[0079] = H + d (1)

[0080] In the formula, H represents the vertical depth of the second-order cut hole; H represents the design cycle advance of the tunnel; and d represents the over-depth value. Considering the strong clamping effect at the bottom of the hole under extremely hard rock conditions, the difficulty in breaking the rock at the bottom of the hole, and the obvious residue at the bottom of the hole, the over-depth value d is taken as 0.2 to 0.4 m to ensure the design cycle advance and improve the utilization rate of the blast hole.

[0081] Based on the cavity formation mechanism of cut-out under extremely hard rock conditions, the vertical depth ratio of the first and second order cut-out holes and the segmented charge ratio of the first and second order cut-out are:

[0082] η= = 0.55~0.65 (2)

[0083] In the formula, , The vertical depths of the first-order and second-order slots are respectively.

[0084] Based on the calculated vertical depth of the second-order cut hole The vertical depth of the first-order slotted hole is calculated using the ratio η. .

[0085] 2. Parameters of slotted perforated mesh

[0086] Center hole: The diameter is the same as the borehole diameter (usually 42mm), and the depth is the same as the second-stage slotting hole;

[0087] First-order slotting holes: inclination angle 65°~70°, hole spacing 0.4~0.5m, row spacing 0.5~0.6m, a total of 8 holes, symmetrically arranged on both sides of the central hole;

[0088] Second-order cut holes: 75°–80° inclination angle, hole spacing 0.5–0.6m, row spacing 0.6–0.7m, a total of 10 holes, arranged surrounding the first-order cut holes, as shown below. Figure 3 As shown;

[0089] Minimum distance between the slotted hole and the auxiliary hole: 0.8~1.0m.

[0090] Through the above-mentioned multi-stage slotting structure design, the initial free space can be gradually formed during the blasting process, reducing the clamping effect under extremely hard rock conditions, and providing a spatial basis for the effective release of explosive energy and the fracturing of deep rock masses under subsequent sealing conditions.

[0091] 3. Initial single-hole charge and segmented charge

[0092] Based on the empirical charge formula modified from Livingston's blasting funnel theory, the total charge per hole for first and second order cut holes is calculated respectively:

[0093] Q max =q·S·L·ɑ (3)

[0094] In the formula, Q maxq represents the total charge per borehole; q represents the empirical value of charge consumption per unit area in extremely hard rock (kg / m³); S represents the bearing area per borehole (m²); L represents the borehole depth (m); and α represents the correction factor for extremely hard rock (≥1.2).

[0095] Among them, based on the cavity formation mechanism of slotting under extremely hard rock conditions, the charge ratio of the first and second order slotting segments is:

[0096] λ= = (4)

[0097] In the formula, This refers to the amount of explosive charged at the orifice. This refers to the amount of explosive charged at the bottom of the borehole.

[0098] The charge amounts at the bottom and the opening of the borehole are respectively:

[0099] = ×40%, = ×60% (5)

[0100] Step 3: Calculate the fill length of the double-sealing system's stemming material.

[0101] To achieve graded energy release of explosives inside the borehole, suppress bursting at the borehole opening, and increase the duration of action of the explosive gases, this invention constructs a dual-sealing system consisting of an "in-hole isolation sealing layer + a borehole-opening composite sealing layer," such as... Figure 4 As shown.

[0102] The borehole isolation and sealing layer is placed between the two sections of explosive charge to prevent the borehole charge from detonating and to achieve the staged release of explosive energy; the borehole opening composite sealing layer is placed on the outside of the borehole to enhance the energy locking effect at the borehole opening and reduce blasting vibration and dust leakage.

[0103] Based on the borehole diameter, borehole depth, and minimum resistance line parameters, determine the length parameters of the in-hole isolation sealing layer and the borehole-mouth composite sealing layer respectively:

[0104] 1. Calculation of the length of the isolation and sealing layer inside the borehole.

[0105] The core function of the internal isolation and sealing layer is to prevent sympathetic detonation and achieve staged release of explosion energy. Its length... Take 8 to 12 times the borehole diameter, and the calculation formula is as follows:

[0106] =(10~12)· (6)

[0107] In the formula: 8 to 10 times for ordinary extremely hard rock scenes, and 10 to 12 times for extremely hard rock scenes; This refers to the diameter of the blast hole; 42mm is commonly used for tunnels.

[0108] This range is matched with the previously designed charge ratio λ. Under this charge ratio, the orifice section and the bottom section need to be physically isolated to prevent sympathetic detonation. 8 to 12 times is a suitable range.

[0109] 2. Calculation of the length of the orifice sealing layer.

[0110] The composite sealing layer at the orifice must simultaneously meet the mandatory requirements of the "Blasting Safety Regulations" and the vibration reduction requirements of extremely hard rock interlocking, with a total sealing length of [missing information]. The following three conditions must be met simultaneously, and the maximum value of the calculation shall be taken: (1) not less than 1 / 3 of the borehole depth; (2) not less than 0.3m; (3) not less than the minimum resistance line W of the borehole.

[0111] 3. Distribution of material length in the composite sealing layer at the orifice.

[0112] like Figure 4 As shown, from the inside out, the layers are: a high-efficiency water-based gun mud layer, a high-density gun mud layer, and a polyurethane foam layer. The length of the high-efficiency water-based gun mud layer is... : Take 1 / 10 to 1 / 5 of the borehole depth; the formula is: water mixed with 0.5% to 1% dish soap / laundry detergent to reduce the surface tension of water and improve the dust suppression and harmful gas adsorption effect; the thickness of the polyurethane foam layer is 25 to 35 cm, completely covering the area around the borehole, and the curing time is ≥15 min; the remaining length is filled with high-density borehole clay in layers and compacted to the total sealing length. .

[0113] Since the initial slotting space has been formed in the previous step, this step reduces the escape of explosive gases and prolongs the action time of the explosive load through a double-sealing structure, so that the explosive energy can be more effectively applied to the formed fractured area, providing an energy basis for the subsequent expansion of the blasting space.

[0114] Step 4: Calculation of Time Delay Parameters for Grading Extremely Hard Rock

[0115] like Figure 5 As shown, based on the initial free space formed by the aforementioned compound cut-out and the energy retention conditions formed by the double-sealing structure, the subsequent borehole detonation sequence and energy release process are rationally controlled according to the evolution process of the blasting space. This is to ensure that subsequent blasting can fully utilize the free surface formed by the preceding blasting, reduce the rock mass fracturing resistance, and achieve the gradual expansion of the fracturing area. This method is based on the high-precision delay capability of electronic detonators and combined with the in-hole segmented blasting rock-breaking mechanism to design a three-level graded detonation sequence.

[0116] Level 1: In-hole graded time delay

[0117] The detonation section at the center hole and the second-order cut hole is pre-charged, with a delay Δ. Then detonate the bottom charge section of the hole. According to the theory of segmented blasting within the hole, Δ It is essential to ensure the stability of the initial free surface formed by the first blasting section and to weaken the clamping effect of the subsequent blasting sections. Based on the engineering practice of segmented blasting in electronic detonator holes in hard rock tunnels, Δ is taken under extremely hard rock conditions. =55~60 ms, where: 55ms for ordinary extremely hard rock (20≤f<30) and 60ms for extremely hard rock (f≥30). Ensure the free surface of the first detonation section is stable before detonating the subsequent section.

[0118] Level 2: Inter-row graded delay

[0119] Detonation is strictly carried out in stages according to the sequence of "first-order cut hole integral charge section and center hole orifice section → second-order cut hole orifice section → center hole and second-order cut hole bottom section", with inter-row staged delay Δ. =50~70ms.

[0120] The process involves three steps: First, detonating the entire charge section of the first-order cut hole and the orifice section of the central hole to create an initial temporary cavity; second, detonating the orifice section of the second-order cut hole to further expand the cavity volume; and third, detonating the central hole and the bottom section of the second-order cut hole to complete deep cutting and rock fracturing at the bottom of the hole. This stepped, staged detonation method creates dynamic free surfaces within the cut cavity, promoting continuous reflection and superposition of subsequent detonation stress waves, while simultaneously weakening the clamping effect of the rock at the bottom of the hole, improving the utilization rate of the explosion energy at the bottom of the hole, thereby enhancing the cutting effect and the cyclic advance.

[0121] Level 3: Same-row staggered peak delay (staggered peak delay Δ) =3~4ms)

[0122] The inter-hole delay between adjacent blast holes in the same auxiliary hole ring is fixed at 4ms, and the inter-hole delay between adjacent blast holes in the same row of peripheral blast holes is fixed at 3ms. This value is based on the principle of peak offset cancellation of electronic detonators in sequential blasting. The blasting time difference between adjacent blast holes should be greater than the minimum delay accuracy of electronic detonators (1ms) and less than the rock fracture penetration time (8-10ms), which can achieve phase offset of vibration waveform and reduce the peak value of blasting vibration.

[0123] All delay parameters are rounded up to an integer multiple of the minimum programmable delay (1ms) of the electronic detonator.

[0124] Step 5: Verification of Single-Stage Explosive Charge

[0125] According to the general Sadovsky empirical formula for predicting the peak vibration velocity of a particle in the "Safety Regulations for Blasting" GB 6722-2014, the formula can be modified to obtain:

[0126] Q max = (7)

[0127] In the formula: Qmax [v] represents the maximum permissible single-stage detonation charge, in kg; [v] represents the safe permissible peak velocity of a mass particle, based on the safety criterion of GB 6722-2014 "Safety Regulations for Blasting", in m / s; R represents the distance from the blast center, in m; K and α are site coefficients, with the empirical range recommended by GB 6722-2014 "Safety Regulations for Blasting" as follows: K = 50~150 under extremely hard rock conditions. =1.3~1.8.

[0128] Using the single-hole charge calculated in Step 2, calculate the total charge for simultaneous detonation of each delayed detonation stage, and compare it with the calculated maximum permissible charge per stage. If the total charge for simultaneous detonation in any delay stage is less than or equal to the maximum permissible charge per stage, the detonation parameter is deemed to meet the blasting vibration safety requirements. If the total charge for simultaneous detonation in any delay stage is greater than the maximum permissible charge per stage, corrections are made by reducing the number of holes detonated simultaneously in the same stage, adjusting the graded delay sequence, or reducing the charge per hole until the blasting vibration safety requirements are met.

[0129] Step 6: Determining the Explosion Effect and Optimizing Parameter Cycle

[0130] After completing the pre-detonation safety inspection, on-site blasting is carried out. The blasting effect is judged through visual indicators, and the blasting parameters are iteratively optimized to determine the final optimized construction plan.

[0131] 1. Core indicators for judging blasting effect: After blasting, four indicators are judged by measuring: ① Hole utilization rate: Measure the length of the remaining hole and calculate the hole utilization rate η=(H-remaining hole length) / H×100%; ② Large block rate: Randomly select three 1m×1m×1m cubic areas in the middle of the blast pile and count the proportion of rock blocks larger than 30cm; ③ Surrounding rock stability and vibration: Observe whether there are falling blocks or cracks in the surrounding rock of the tunnel and measure the stability of the surrounding rock; ④ Sealing effect: Observe whether there is any blasting at the borehole or leakage of blast gas.

[0132] 2. Parameter optimization rules for non-compliance: ① If borehole utilization rate < 93%: Increase the proportion of the bottom charge section or extend the segment delay Δ within the borehole. (Maximum 60ms); ② If the large block ratio is >8%: shorten the delay Δ between the slotted hole and the auxiliary hole row. (Lower limit 50ms), or increase the proportion of the charging section at the orifice (upper limit 45%); ③ If rockfall occurs and the surrounding area experiences significant tremors: increase the inter-row delay Δ (Maximum 70ms), or reduce the maximum amount of explosive charge per stage; ④ If blasting or excessive dust occurs at the orifice: increase the total length of the composite sealing layer at the orifice. Alternatively, the polyurethane foam sealing layer can be thickened.

[0133] 3. Iterative Cyclic Rules: The optimized blasting parameters are used as the initial input parameters for the next blasting cycle, and the design and verification process from Step 2 to Step 5 is repeated. When two consecutive blasting cycles meet the design requirements of "borehole utilization rate ≥ 93%, large block rate ≤ 8%, surrounding rock stability meets the standard, and no impact blasting", the iterative cyclical process is terminated, and the final optimized construction parameters for blasting of extremely hard rock tunnels are determined.

[0134] This invention improves the utilization rate of explosive energy and the precise control of blasting vibration by optimizing the compound wedge-shaped slotting structure, designing a double-seal segmented charging system, setting up a graded precise delay detonation network, and establishing a low-equipment-dependent visual parameter closed-loop optimization method. It solves the technical defects of existing technologies, such as poor slotting effect in extremely hard rock tunnels, low borehole utilization, easy exceedance of blasting vibration velocity, and high threshold for on-site construction, and achieves the unity of efficient rock breaking and micro-vibration control.

[0135] The invention will be further illustrated below with an example of a domestic extremely hard rock tunnel project.

[0136] Step 1: Determine the parameters of the extremely hard rock mass

[0137] On-site sampling and testing of the rock mass at the working face yielded the following basic rock parameters: rock density. =2.75 g / Uniaxial compressive strength of rock =230 MPa, Protodyakonov hardness coefficient of rock f=23 (classified as ordinary extremely hard rock), rock P-wave velocity =5200 m / s.

[0138] Step 2: Calculate the key parameters of the slotting hole

[0139] A three-stage compound wedge arrangement of "center hole + first-stage slotting hole + second-stage slotting hole" is adopted, and the diameter of the blast hole is D=42mm.

[0140] 1. Structural Parameter Calculation: In this embodiment, the designed cyclic advance is H=2.7m. Considering the bottom clamping effect under extremely hard rock conditions with f=23, the over-depth value is taken as d=0.3m. Therefore, the vertical depth of the second-order cut hole is:

[0141] =H+d=2.7+0.3=3.0m;

[0142] Based on the parameters of extremely hard rock, and taking the first-order and second-order depth ratio η=0.6, the vertical depth of the first-order cut hole is calculated. =1.8m.

[0143] 2. Hole Network Parameter Layout: Three 42mm central holes with a depth of 3.0m are arranged in the center, using a segmented charging structure with the hole opening and bottom sections, and an internal isolation and sealing layer is set between the two sections to achieve staged delayed initiation in the hole; four first-order cut holes (68° inclination angle, hole spacing 0.45m) are arranged on one side, for a total of eight on both sides; five second-order cut holes (78° inclination angle, hole spacing 0.5m) are arranged on one outer side, for a total of ten on both sides; the row spacing between the first-order cut holes and the second-order cut holes is 0.6m; the minimum distance between the second-order cut holes and the surrounding auxiliary holes is maintained at 0.9m.

[0144] 3. Initial single-hole charge and segmented charge calculation: The empirical estimate of the unit charge consumption for extremely hard rock is q = 2.0 kg / Substituting into the modified empirical charge formula Q = q·S·L·ɑ (here, the orifice loading area S is set to 0.4), we arrive at the corrected formula. (Hole depth L = 3.0m, correction factor α = 1.3), calculate the initial single-hole charge Q for the center hole and the second-order cut hole: Q = 2.0 × 0.4 × 3.0 × 1.3 = 3.12 kg. Single-hole opening charge. =3.12 × 40% = 1.248 kg, amount of medicine at the bottom of the hole =3.12 × 60% = 1.872 kg. First-order cut hole depth = ×0.6=1.8m, substituting into the modified Livingston formula, we obtain the initial single-hole charge for the first-order cut hole. =2.0×0.4×1.8×1.3 = 1.872kg.

[0145] Step 3: Calculate the fill length of the double-sealing system's stemming material.

[0146] Based on calculations, segmented charging and dual-system plugging were performed on the central hole and the second-order deep-cut hole, respectively:

[0147] 1. In-hole isolation and sealing layer: Based on extremely hard rock with f=23, a layer of 10 times the borehole diameter is used, i.e. = 10×42mm=420mm=0.42m.

[0148] 2. Orifice composite sealing layer: Total orifice sealing length = max(3.0 / 3, 0.3m, W=0.9m)=1.0m.

[0149] 3. Material layer distribution: From the inside of the hole to the outside, first fill with 0.4m of high-efficiency water-based drilling mud; then compact with about 0.3m of high-density traditional drilling mud; finally, pour in a polyurethane foam layer with a thickness of about 0.3m and wait for it to cure for more than 15 minutes.

[0150] Step 4: Calculation of time delay parameters for extremely hard rock grading (selecting electronic detonator initiation network)

[0151] Design a precise time-delayed detonation sequence for three levels of extremely hard rock:

[0152] 1. First stage (in-hole staged delay): Both the center hole and the second-stage slotted hole adopt an in-hole staged delay detonation structure. The hole opening section detonates first, and the bottom section detonates after a delay of 55ms.

[0153] 2. Second level (inter-row graded delay): Detonation is strictly in sequence, with an inter-row delay Δ The detonation time was set to 60ms (with no sensitive buildings nearby). The actual detonation sequence was as follows: the entire charge section of the first-stage cut hole and the opening section of the central hole were detonated simultaneously; after a 60ms delay, the opening section of the second-stage cut hole was detonated; after a further 55ms delay within the hole, the bottom section of the central hole and the bottom section of the second-stage cut hole were detonated; then, after another 60ms delay, the outer auxiliary holes and surrounding holes were detonated.

[0154] 3. Level 3 (Same Row Staggered Delay): The delay between adjacent blast holes in the same row of auxiliary holes is set to 4ms, and the delay between peripheral blast holes is set to 3ms.

[0155] Step 5: Verification of Single-Stage Explosive Charge

[0156] According to the "Safety Regulations for Blasting", the safe permissible vibration velocity for this condition is [v] = 3.0 cm / s. Based on the Sadovsky deformation formula... Taking the extremely hard rock site coefficient K=100 and α=1.5, the maximum charge at a distance R=40m from the blast center is calculated to be 27kg.

[0157] The charge ratio is λ=60%, meaning that the charge at the bottom of the center hole and the second-order slotted hole accounts for 60% of the total charge in a single hole, while the charge at the opening accounts for 40%.

[0158] First stage detonation:

[0159] The first-order cut holes and the central hole opening section were detonated simultaneously. There were a total of 8 first-order cut holes.

[0160] =8 × 1.872 = 14.976 kg;

[0161] There are 3 central orifice sections, with a total charge per orifice of 3.12 kg, and the orifice sections account for 40% of the total charge.

[0162] =3 × 3.12 × 40% = 3.75 kg;

[0163] First stage total dosage:

[0164] = + =18.726kg < 27kg;

[0165] The amount of explosive used in the first stage of detonation meets safety requirements.

[0166] Second stage detonation:

[0167] The second-stage cut-out holes are detonated at the orifice section. There are a total of 10 second-stage cut-out holes, with a total charge of 3.12 kg per hole. The charge at the orifice section accounts for 40% of the total charge. The total charge in the second stage is:

[0168] =10×3.12×40%=12.48kg<27kg

[0169] The amount of explosive used in the second stage of detonation meets safety requirements.

[0170] Third-stage detonation:

[0171] The bottom sections of the central holes and the second-stage cut holes are detonated simultaneously. There are three central holes in total, with the bottom sections accounting for 60% of the charge.

[0172] =3 × 3.12 × 60% = 5.62 kg

[0173] The bottom section of the second-order cut holes accounts for 60% of the total number of holes. There are a total of 10 second-order cut holes.

[0174] =10×3.12×60%=18.72kg

[0175] Total dosage in the third stage

[0176] = + =24.34kg<27kg

[0177] The amount of explosive used in the third stage of detonation meets safety requirements.

[0178] Therefore, it meets the requirements.

[0179] Step 6: Determine the final blasting parameters by determining the cycle.

[0180] 1. Initial blasting implementation and judgment: The blasting was carried out according to the above parameters. After the blast, the following measurements and tests were conducted: the average length of the residual hole was 0.3m, the hole utilization rate η = (3.0-0.3) / 3.0 = 90% (less than the standard of 93%); there was no obvious blasting phenomenon; the large block rate was about 6% (meeting the standard).

[0181] 2. Parameter Optimization and Iteration: Triggering the non-compliance optimization rule "If borehole utilization rate < 93%, increase the upper limit of the segmented delay within the borehole". Adjust the segmented delay Δ within the borehole for the next cycle. The maximum value has been adjusted from 55ms to 60ms.

[0182] 3. Secondary blasting implementation: Adjusted parameters were used for drilling and blasting. The average length of residual holes measured after blasting decreased to 0.15m, and the borehole utilization rate increased to 95%; the large block rate decreased to 5%, and there was no abnormal dust ejection from the borehole opening, indicating stable surrounding rock at the tunnel face. After two consecutive cycles of achieving the required indicators and no longer decreasing with further advance, iteration was stopped, and this set of parameters was solidified as a standardized blasting scheme for this extremely hard rock tunnel.

[0183] To verify the synergistic effect among Step 2 (compound wedge-shaped slotting structure), Step 3 (double sealing system), and Step 4 (three-level delay system) of this invention, comparative calculations were performed under the same extremely hard rock tunnel conditions (Protodyakonov coefficient f=23, uniaxial compressive strength 230MPa).

[0184] The same Step 2 slotting parameters and Step 4 delay parameters as in this invention are used, but the in-hole isolation sealing layer and the orifice composite sealing layer are omitted, and only conventional stemming material is used for sealing. Based on the positive correlation between the pressure holding time of explosive gas and rock breaking efficiency, it is estimated that the high in-hole pressure generated by blasting of extremely hard rock will force open the orifice under conventional stemming material conditions, causing the explosive gas to dissipate prematurely. During the inter-row delay period set in Step 4, there is no gas expansion work to assist rock breaking, resulting in a lack of beneficial effects from continuous gas work when subsequent rows of blast holes are blasted. It is estimated that the blast hole utilization rate is lower than that of the complete scheme, and the phenomenon of blasting at the orifice occurs.

[0185] The same Step 2 slotting parameters and Step 3 double-sealing system as the present invention are adopted, but the in-hole staged delay and inter-row staged delay are eliminated (each section is detonated simultaneously). Based on the principle of vibration wave superposition, it is estimated that the simultaneous detonation of each borehole leads to the superposition of blasting vibration waves in the same direction, and the peak vibration velocity is about 30% higher than that estimated by the present invention; at the same time, the subsequent boreholes are detonated before the free surface of the preceding boreholes has been formed, and the rock-breaking energy of each section of charge cancels each other out. It is estimated that the borehole utilization rate is lower than that of the complete scheme, and the blasting vibration velocity exceeds the safety threshold.

[0186] This invention employs the same Step 3 double-sealing system and Step 4 delay parameters as the present invention, but uses a conventional single-stage slotting method. Based on the principle of progressive expansion of the slotted cavity, it is deduced that due to the lack of a multi-stage slotting space formation process, the ability to form a free surface in the early stage of blasting is reduced, and the rock mass clamping effect is enhanced, resulting in insufficient usable space in subsequent blasting processes, and the sealing structure and graded delay measures cannot be fully utilized. The calculated borehole utilization rate is lower than the complete scheme, and the slotted cavity volume is significantly insufficient.

[0187] The above comparison shows that there is a significant synergistic effect among Steps 2, 3, and 4. The absence of any one of these elements prevents the achievement of the comprehensive blasting effect of this invention, which features "hole utilization rate ≥ 93%, large block rate ≤ 8%, and no-impact blasting." This demonstrates that the three technical features of this invention, as a whole, produce a non-obvious synergistic effect.

[0188] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any way. Although the present invention has been disclosed above with reference to preferred embodiments, it is not intended to limit the present invention. Any person skilled in the art can make some modifications or alterations to the above-disclosed technical content to create equivalent embodiments without departing from the scope of the present invention. Any simple modifications, equivalent changes, and alterations made to the above embodiments based on the technical essence of the present invention without departing from the scope of the present invention shall still fall within the scope of the present invention.

Claims

1. An optimized method for staged vibration reduction blasting in the complex wedge-shaped cut-out double-sealing stage in extremely hard rock tunnels, characterized in that... Includes the following steps: Step 1: Determine the parameters of the extremely hard rock mass On-site samples were taken from the rock mass to be blasted to measure its basic physical and mechanical parameters, including rock density. Rock longitudinal wave velocity Rock static Poisson's ratio Uniaxial compressive strength of rock Rock elastic modulus Rock Protodextrin coefficient ƒ and rock tensile strength ; Step 2: Calculate the key parameters of the slotting hole The tunnel adopts a three-stage compound wedge-shaped arrangement of "central hole + first-stage cut hole + second-stage cut hole". There are 3 central holes, which are arranged vertically along the tunnel's central axis. There are 8 first-stage cut holes and 10 second-stage cut holes. Determine the core parameters using the following formula: (1) Vertical depth of the second-order slotted hole: = H + d, where H is the tunnel design cycle advance, and d is the over-depth value, d=0.2~0.4m; (2) Vertical depth of first-order slotted hole: H1 = η·H2, where η = H1 / H2 = 0.55~0.65; (3) Total charge per hole: Q = q·S·L·ɑ, where q is the empirical value of charge per unit in extremely hard rock, S is the bearing area of ​​a single hole, L is the depth of the hole, and ɑ is the correction coefficient for extremely hard rock. (4) Segmented charge ratio: λ=Q1 / Q2=2 / 3, where Q1 is the charge amount at the orifice opening and Q2 is the charge amount at the orifice bottom, and Q1=Q×40% and Q2=Q×60%; Step 3: Calculate the fill length of the double-sealing system's stemming material. Construct a dual-sealing system consisting of an "in-hole isolation sealing layer + orifice composite sealing layer," with the in-hole isolation sealing layer having a length of [missing information]. =(8~12)· ,in The borehole diameter; the total length of the composite sealing layer at the borehole opening. The following conditions must be met: (1) Not less than 1 / 3 of the borehole depth; (2) Not less than 0.3m; (3) Not less than the minimum resistance line W of the borehole. The orifice composite sealing layer consists of, from the inside out, a high-efficiency water gun mud layer, a high-density gun mud layer, and a polyurethane foam layer. Step 4: Calculation of Time Delay Parameters for Grading Extremely Hard Rock Design a three-stage detonation sequence: Level 1: In-hole graded delay, the charging section at the detonation orifice in the center hole and the second-stage cut-out hole has a delay Δ. Then, the explosive charge section at the bottom of the detonation hole; Level 2: Inter-row graded delay, the detonation sequence is: the first-order slotted hole integral charge section and the center hole opening section → the second-order slotted hole opening section → the center hole and the bottom section of the second-order slotted hole. Level 3: Staggered detonation in the same row, inter-hole detonation of adjacent blast holes in the same ring of auxiliary holes, and inter-hole detonation of adjacent blast holes in the same row of peripheral blast holes. Step 5: Verification of Single-Stage Explosive Charge The maximum permissible amount of explosive charge per stage is calculated based on the Sadovsky empirical formula. The total amount of explosive charge for all stages detonated simultaneously, calculated in Step 2, is compared with the maximum permissible amount of explosive charge per stage. If the requirements are not met, adjustments are made. Step 6: Determining the Explosion Effect and Optimizing Parameter Cycle The blasting effect is determined by visual indicators, and the blasting parameters are iteratively optimized until two consecutive blasting cycles meet the design requirements.

2. The optimized method for compound wedge-shaped slotting double-sealing graded vibration reduction blasting in extremely hard rock tunnels according to claim 1, characterized in that, The mesh parameters for the slotted holes in Step 2 are as follows: The diameter of the center hole is the same as the diameter of the blast hole, and the depth is the same as that of the second-order slotting hole; The inclination angle of the first-stage slotting holes is 65°~70°, the hole spacing is 0.4~0.5m, and the row spacing is 0.5~0.6m; The inclination angle of the second-order slotting holes is 75°~80°, the hole spacing is 0.5~0.6m, and the row spacing is 0.6~0.7m; The minimum distance between the slotted hole and the auxiliary hole is 0.8~1.0m.

3. The optimized method for compound wedge-shaped slotting double-sealing staged vibration reduction blasting in extremely hard rock tunnels according to claim 1, characterized in that, In Step 3: The length of the isolation and sealing layer inside the borehole is 8 to 10 times the borehole diameter in ordinary extremely hard rock scenarios, and 10 to 12 times the borehole diameter in extremely hard rock scenarios. The length of the mud layer in the high-efficiency water cannon is 1 / 10 to 1 / 5 of the borehole depth. Its formula is water and 0.5% to 1% surfactant, wherein the surfactant is detergent or laundry powder. The thickness of the polyurethane foam layer is 25~35cm, and the curing time is ≥15min; High-density blasting mud layer is compacted and filled in layers to the total sealing length. .

4. The optimized method for compound wedge-shaped slotting double-sealing graded vibration reduction blasting in extremely hard rock tunnels according to claim 1, characterized in that, In Step 4: the first-stage in-hole graded delay Δ The second-level inter-row delay Δ is 55-60ms. The delay time is 50~70ms, and the second-level staggered peak delay is 3~4ms. All delay parameters are rounded up to an integer multiple of the minimum programmable delay of the electronic detonator.

5. The optimized method for compound wedge-shaped slotting double-sealing staged vibration reduction blasting in extremely hard rock tunnels according to claim 1, characterized in that: In Step 5, the formula for predicting the peak vibration velocity of a particle, based on the general Sadovsky empirical formula in the "Safety Regulations for Blasting" GB 6722-2014, can be modified to obtain: Q max = ; In the formula: Q max [v] is the maximum permissible single-stage detonation charge; [v] is the safe permissible peak particle vibration velocity; R is the distance from the detonation center; K and α are site coefficients, with K = 50–150 under extremely hard rock conditions. =1.3~1.8; When the total charge of explosives detonated simultaneously during any level of delay is less than or equal to the maximum permissible charge of explosives in a single stage, the detonation parameter is deemed to meet the blasting vibration safety requirements. When the total charge of explosives detonated simultaneously during any level of delay is greater than the maximum permissible charge of explosives in a single stage, the method of correction is to reduce the number of blast holes detonated simultaneously in the same stage, adjust the order of graded delays, or reduce the charge of explosives in a single hole until the blasting vibration safety requirements are met.

6. The optimized method for graded vibration reduction blasting of double-sealed, double-wedge-cut, and multi-stage rock tunnels according to claim 1, wherein the uniaxial compressive strength of the extremely hard rock mass is ≥200MPa and the Protodyakonov robustness coefficient f is ≥20.

7. The optimized method for graded vibration reduction blasting in double-seal, compound wedge-shaped cut-out tunnels in extremely hard rock according to any one of claims 1-6, wherein in Step 6, the core evaluation index for blasting effect includes: ① Hole utilization rate: Measure the remaining length of the blast holes and calculate the blast hole utilization rate η = (H - remaining hole length) / H × 100%; ② Large block rate: Randomly select three 1m × 1m × 1m cubic areas in the middle of the blast pile and count the proportion of rock blocks larger than 30cm; ③ Surrounding rock stability and vibration: Observe whether there are falling blocks or cracks in the surrounding rock of the tunnel and determine the stability of the surrounding rock; ④ Sealing effect: Observe whether there is any blasting at the borehole or leakage of blast gas; The parameter optimization rules for cases that do not meet the standards are as follows: ① If the borehole utilization rate is <93%: increase the proportion of the charge section at the bottom of the borehole, or extend the segmented delay Δ in the borehole. ② If the large block ratio is >8%, shorten the time delay Δ between the slotted hole and the auxiliary hole row. ③ If rockfall occurs and there is significant tremor in the surrounding area: increase the inter-row delay Δ ④ If blasting or excessive dust occurs at the orifice: increase the total length of the composite sealing layer at the orifice. Alternatively, the polyurethane foam sealing layer can be thickened.

8. The optimized method for compound wedge-shaped cut-out double-sealing graded vibration reduction blasting in extremely hard rock tunnels according to claim 7, characterized in that, The termination condition for the iterative cycle in Step 6 is: two consecutive blasting cycles must meet the design requirements of borehole utilization rate ≥93%, large block rate ≤8%, surrounding rock stability, and no impact blasting.

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