Blast control construction method for station in rock stratum zone by using arch-cover method

Abstract

A blast control construction method for a station in a rock stratum zone by using an arch-cover method. The blast control construction method comprises the following steps: S1, determining an excavation solution on the basis of a construction blueprint, and dividing an entire station section into an arch-cover portion and a straight-wall portion; S2, determining drilling positions of peripheral holes in each block; S3, determining the drilling positions of cut holes in each block at an arch portion; S4, determining drilling positions of auxiliary holes in each block; S5, determining drilling positions of floor holes in each block; S6, performing drilling according to the confirmed drilling positions; and S7, after the drilling is completed, performing charging according to the solution, sequentially performing detonation according to the excavation solution and the position of each detonator section, optimizing the charged amount and blasthole design for blasting of lower rocks, and reserving temporary support rock mass. The method can reduce the impact of blast vibration of a lower rock stratum on surrounding rocks of arch feet, reduce explosives consumption, and ensure the safety of the temporary arch supports.

Description

Rock stratum arch cover method station blasting control construction method TECHNICAL FIELD

[0001] The present application relates to the technical field of underground engineering construction, and in particular to a rock stratum arch cover method station blasting control construction method. BACKGROUND

[0002] In the rock stratum city, the arch cover method is a unique advantage of the underground excavation method. By excavating the upper half of the stratum with poor geological conditions and directly applying the arch cover secondary lining, the stability of the lower rock mass can be effectively utilized, the arch cover settlement can be controlled, and the high-strength secondary lining structure can effectively prevent the deformation of the "upper soft". When the lower rock mass is blasted, the arch cover and the working space below the arch cover can also play a role in shock absorption and safety protection.

[0003] At present, the arch cover method has been widely used in the construction of subways in rock stratum cities, but the specific blasting construction design scheme mainly depends on past experience and on-site test judgment, and there is a lack of systematic blasting construction design method for the characteristics and process requirements of the arch cover method.

[0004] Therefore, the designer of the present application, in view of the above-mentioned defects, through intensive research and design, and based on years of experience and achievements in the relevant industry, improves the original construction method according to the characteristics and process requirements of the arch cover method, and designs a rock stratum arch cover method station blasting construction method to solve the above-mentioned problems. This method has the advantages of small amount of explosives, good arch foot protection effect, and small construction vibration, and is an economical, safe and green blasting construction method.

[0005] SUMMARY

[0006] The purpose of the present application is to provide a rock stratum arch cover method station blasting control construction method, which divides the whole section of the arch cover method into upper and lower parts for excavation, optimizes the blasting timing and arch foot protection measures according to the requirements of the arch cover method for arch foot protection, and effectively reduces the influence of lower rock stratum blasting vibration on the arch foot surrounding rock. Based on the characteristics of the arch cover method of excavating the upper half of the stratum with poor geological conditions and directly applying the arch cover secondary lining to create a new free surface, the charge amount and blast hole design for the lower rock blasting are optimized, the amount of explosives is reduced, and the construction cost is reduced. In order to improve the stability of the arch excavation process, the temporary support rock mass is reserved to ensure the safety of the temporary arch.

[0007] To achieve the above-mentioned purpose, the present application discloses a rock stratum arch cover method station blasting control construction method, characterized in that it comprises the following steps:

[0008] S1. The station section is divided into an arch cover part and a straight wall part, wherein the arch cover part includes three parts and is excavated by using a double side wall pilot tunnel method, the straight wall part includes nine parts and is excavated by using a slope excavation method, so that a total of 12 parts are excavated;

[0009] S2. The positions of the peripheral hole drilling of each sub-block are determined;

[0010] S3. The positions of the slotting hole drilling of each sub-block of the arch part are determined;

[0011] S4. The positions of the auxiliary hole drilling of each sub-block are determined;

[0012] S5. The positions of the floor hole drilling of each sub-block are determined, and the floor hole positions are determined according to the rock area that needs to be blasted on the section;

[0013] S6. According to the determined drilling positions, the peripheral hole, the auxiliary hole, the slotting hole and the floor hole are drilled;

[0014] S7. After the drilling is completed, the charging is carried out according to the scheme, and the detonation is carried out according to the excavation scheme and the positions of each detonator section in turn.

[0015] The step S1 specifically includes the following sub-steps:

[0016] S11. The upper advanced small guide pipe of the right side pilot tunnel of the arch cover part is constructed, the right side pilot tunnel I of the arch cover part is excavated, and the initial support of the right side pilot tunnel of the arch cover part is constructed;

[0017] S12. The upper advanced small guide pipe of the left side pilot tunnel of the arch cover part is constructed, the left side pilot tunnel of the arch cover part is excavated when the right side pilot tunnel of the arch cover part is advanced by not less than 15m than the left side pilot tunnel, and the initial support of the left side pilot tunnel of the arch cover part is constructed;

[0018] S13. The upper advanced small guide pipe of the middle pilot tunnel of the arch cover part is constructed, the middle pilot tunnel of the arch cover part is excavated by using a bench method when the left side pilot tunnel of the arch cover part is advanced by not less than 15m than the middle pilot tunnel of the arch cover part;

[0019] S14. After the excavation and support of the right side pilot tunnel, the left side pilot tunnel and the middle pilot tunnel of the arch cover part are completed, the middle temporary cross brace is removed in sections, and the structural arch cover lining is constructed;

[0020] S15. The left part of the middle upper pilot tunnel of the straight wall part is excavated, the right upper pilot tunnel of the straight wall part is constructed again when the left part of the middle upper pilot tunnel is advanced by 5m, and the support is carried out;

[0021] S16. The right part of the middle upper pilot tunnel of the wall part is excavated, the left upper pilot tunnel is constructed again when the right part of the middle upper pilot tunnel is advanced by 5m, and the support is carried out; S17. The middle upper pilot tunnel is excavated 4-5m after the middle middle pilot tunnel;

[0022] S18. After the middle middle pilot tunnel advances 5m to the right middle pilot tunnel and the left middle pilot tunnel, excavate the right middle pilot tunnel and the left middle pilot tunnel and support them;

[0023] S19. After the middle middle pilot tunnel advances 4-5m to the middle lower pilot tunnel, excavate the middle lower pilot tunnel, and the middle lower pilot tunnel advances 4.0m per cycle;

[0024] S110. After the middle lower pilot tunnel advances 5-10m to the right lower pilot tunnel and the left lower pilot tunnel, excavate the right lower pilot tunnel and the left lower pilot tunnel and support them.

[0025] Wherein: in the step S11, S12, the initial support construction is carried out for the arch part two side pilot tunnels, at the same time, 50cm thick side wall is excavated in advance from the upper arch cover part to the lower part which is not excavated, so as to advance the initial support for the side wall part, the side wall below uses tunnel spoil layered ramming compaction to form multi-layered spoil compact layer, and C15 concrete layer is constructed on the top surface of the spoil compact layer.

[0026] Wherein: in the step S15, S16, the right upper pilot tunnel and the left upper pilot tunnel are constructed, the spoil compact layer and the C15 concrete layer are cleaned in advance, and the iron sheet layer is used to cover the waterproof roll material, and the outer corner of the iron sheet layer is wrapped with angle steel.

[0027] Wherein: the step S2 comprises the following steps:

[0028] S21. It is necessary to determine whether to use smooth blasting scheme or pre-splitting blasting, smooth blasting is used for the geological conditions of high weathering degree and broken surrounding rock, and pre-splitting blasting is used for the geological conditions of hard and complete surrounding rock;

[0029] S22. The peripheral hole interval is generally 40-50cm for soft rock, 45-55cm for medium-hard rock, and 50-60cm for hard rock;

[0030] S23. In order to reduce the disturbance to the surrounding rock, holes are drilled between the peripheral holes as damping holes;

[0031] S24. Because the arch part is excavated by using double side wall pilot tunnel method, in order to avoid damage to the two sides of the temporary arch, two times of blasting are carried out, 70cm rock body close to the steel frame on both sides is reserved for secondary excavation, and holes are drilled according to the peripheral hole design interval in the temporary support rock body close to the middle during construction, but no charge is carried out, after the completion of the first blasting, the 70cm rock body is basically loose, and mechanical excavation can be carried out, if it is not loose, secondary charge is carried out for supplement, and then mechanical excavation is carried out.

[0032] Wherein: the step S3 comprises the following steps:

[0033] S31. Using wedge-shaped cutting blasting in the throw blasting, the wedge-shaped cutting cutting holes are arranged in pairs in the vertical direction, and a wedge-shaped slot is blasted after blasting;

[0034] S32. The height h of the cutting hole from the floor is 0.5m-0.8m;

[0035] S33. The cutting hole is 0.2-0.3m deeper than the auxiliary hole, and the design parameters are shown in Table 1;

[0036] Table 1 Cutting parameter table

[0037] Wherein a is the vertical distance between the cutting holes, B is the horizontal distance between the cutting holes, and γ is the angle between the hole direction on the plane and the working face.

[0038] Wherein: the step S4 comprises the following steps:

[0039] S41. The auxiliary hole spacing is 50-90cm, and the control resistance line is 1.05-1.25 times the peripheral hole spacing, with the large value for hard rock and the small value for soft rock;

[0040] S42. The overall principle of the hole arrangement is to reduce the number of holes as much as possible, and the arrangement of the auxiliary hole should be carried out after the arrangement of the cutting hole and the peripheral hole, and the number of holes per unit area should be between 1.5-4.5 / m2, and the specific calculation is carried out according to the formula Wherein N is the number of holes, q is the explosive unit consumption, S is the area of the auxiliary hole excavation range, r is the linear charge density of the explosive (kg / m), and n is the hole charge coefficient.

[0041] Wherein: the step S6 comprises the following steps:

[0042] S61. The auxiliary hole depth is determined according to the cycle footage, and each cycle footage is a grid steel frame, and the bottom of the hole is 0.2m-0.3m longer than the hole;

[0043] S62. The peripheral hole and the floor hole are 0-0.2m deeper than the auxiliary hole, and the cutting hole is 0.1-0.3m deeper than the auxiliary hole;

[0044] S63. According to the positions of the peripheral hole, the auxiliary hole, the cutting hole and the floor hole determined by S1-S5, and according to the peripheral hole, the auxiliary hole, the cutting hole and the floor hole depth determined by S61 and S62, drilling is carried out.

[0045] Wherein: the step S7 comprises the following steps:

[0046] S71. Calculate the single-hole charge and charge, determine the single-hole charge of different holes, and the calculation formula is Wherein N is the number of blast holes, q is the explosive unit consumption, S is the auxiliary hole excavation range area, r is the linear charge density of explosive (kg / m), n is the blast hole charge coefficient, and L is the blast hole length.

[0047] S72. Single-stage simultaneous maximum charge calculation is performed according to the formula in section 13.2.4 of the Blasting Safety Regulations (GB6722-2014) Verification is performed, wherein: K and a are a coefficient and an attenuation coefficient related to the terrain and geology between the blasting point and the blasting object, generally, K is 200, and a is 1.8; Q is the maximum charge of single-stage simultaneous blasting, in kg; R is the distance from the blasting center point to the protected object, in m; and V is the calculated blasting vibration speed, in cm / s.

[0048] S73. Delayed blasting interference vibration reduction technology is adopted to reduce the maximum blasting vibration speed and the maximum detonating charge.

[0049] S74. After checking that the blasting network is correct, initiation is performed.

[0050] It can be known from the above content that the station blasting control construction method of the rock stratum arch cover method has the following effects: 1. The entire station section is divided into two large parts and 12 small parts for excavation, wherein the arch cover part includes three parts, the straight wall part includes nine parts, and the arch cover part is excavated in a double-side-wall pilot tunnel mode, and the straight wall part is excavated in a slope mode. After the upper arch cover part is excavated, the arch cover second lining is constructed, and then the lower straight wall part is excavated. The method can utilize the stability of the lower rock mass, control the arch cover settlement, the high-strength second lining structure can effectively prevent the deformation of the "upper soft", and the arch cover and the working space below the arch cover can also play a role in shock absorption and safety protection when the lower rock mass is blasted. In addition, the slope excavation of the lower straight wall part does not need the throwing blasting of the traditional tunnel section excavation, does not need slotting, and only needs to be loosened and blasted, so that the explosive unit consumption can be greatly reduced, and the economic benefit can be improved.

[0051] 2. When the arch part two-side pilot tunnels I and II are constructed, the side wall is excavated by 50 cm downward in advance, the side wall part is preliminarily supported in advance, the excavated part is densely filled by using tunnel spoil (the maximum particle size is not greater than 10 cm) in layers, the thickness of each layer is 10 cm, a 10 cm thick C15 concrete is constructed on the top surface, the rubble and the C15 concrete are cleaned in advance when the lower V and VI parts are blasted and constructed in steps S15 and S16, the waterproof material is wrapped by using 1.5 mm thick iron sheet, the external edges are wrapped by using L56*36*3 angle steel, the outer side is blasted first, and then the side wall part is blasted, which can effectively reduce the damage of blasting to the arch cover structure and the waterproof roll material, and improve the construction quality.

[0052] 3. Because the arch section is excavated using the double-sided wall pilot tunnel method, to avoid damage to the temporary arch frames already erected on both sides, blasting is carried out in two stages. A 70cm section of rock adjacent to the steel frame is left for secondary excavation. During construction, holes are drilled in the rock mass near the central temporary support according to the designed spacing of the surrounding holes, but no explosives are loaded. After the first blast, the 70cm rock mass is basically loosened, and mechanical excavation can be used. If it is not loosened, a second blast can be carried out, followed by mechanical excavation. This method can effectively reduce the impact of construction in the central Zone III on the temporary arch frames already erected in the pilot tunnels on both sides, improving the safety of the arch cover method construction.

[0053] 4. Based on the characteristics and technological requirements of the arch-type blasting method, the original construction method was improved, and a blasting design method suitable for the arch-type blasting method was proposed. The explosive consumption per unit area was optimized to save costs, taking into account the characteristics of upper-part throwing blasting and lower-part loosening blasting in the arch-type blasting method. Corresponding protective measures were proposed to address the characteristics of the arch structure's reverse construction, which can improve the quality of structural construction. To address the risk of damage and failure of the temporary steel frame during the arch excavation process, the requirement of retaining the supporting rock mass for secondary excavation was proposed. This achieved economical, safe, and environmentally friendly cross-sectional blasting construction operations for the arch-type blasting method.

[0054] The details of this invention will become apparent from the following description and the accompanying drawings. Attached Figure Description

[0055] Figure 1 shows a schematic diagram of the construction process of the rock stratum arch-type railway station blasting control method of the present invention.

[0056] Figure 2A shows a schematic diagram of the arch foot protection measures during the construction of the upper arch cover in this invention.

[0057] Figure 2B shows a schematic diagram of the arch foot protection measures during the lower part of the construction in this invention.

[0058] Figure 3A shows a schematic diagram of drilling at the working face in this invention.

[0059] Figure 3B shows a cross-sectional view along the MM direction in Figure 3A.

[0060] Figure 4 shows a schematic diagram of the cross-sectional borehole arrangement in this invention.

[0061] Figure 5 shows a schematic diagram of a specific embodiment of the present invention. Detailed Implementation

[0062] Referring to Figures 1 to 4, the rock formation arch cover method for controlling the blasting construction of railway stations according to the present invention is shown.

[0063] The method for controlling the blasting construction of the rock strata arch-type railway station includes the following steps:

[0064] S1. Referring to Fig. 1, according to the construction blueprint provided by the design, a scheme of excavation is determined, the entire station section is divided into an arch cover part and a straight wall part, the arch cover part includes three parts, the arch cover part is excavated by using a double side wall pilot tunnel method, the straight wall part includes nine parts, the straight wall part is excavated by using a slope method, and thus the entire station section is divided into twelve parts in total for excavation.

[0065] Preferably, the step S1 specifically includes the following sub-steps:

[0066] S11. An upper advanced small guide pipe of a right side pilot tunnel I of the arch cover part is constructed, the right side pilot tunnel I of the arch cover part is excavated, initial support of the right side pilot tunnel I of the arch cover part is constructed, and each cycle is 0.5 m;

[0067] S12. An upper advanced small guide pipe of a left side pilot tunnel II of the arch cover part is constructed, the left side pilot tunnel II of the arch cover part is excavated when the right side pilot tunnel I of the arch cover part is advanced by not less than 15 m than the left side pilot tunnel II, initial support of the left side pilot tunnel II of the arch cover part is constructed, and each cycle is 0.5 m;

[0068] S13. An upper advanced small guide pipe of a middle pilot tunnel III of the arch cover part is constructed, the middle pilot tunnel III of the arch cover part is excavated by using a step method when the left side pilot tunnel II of the arch cover part is advanced by not less than 15 m than the middle pilot tunnel III of the arch cover part, and more specifically, an upper step III-1 is excavated first, initial support of the upper step III-1 is constructed, a lower step III-2 is excavated and supported when the upper step III-1 is advanced by about 5 m than the lower step III-2, and each cycle is 0.5 m;

[0069] S14. After the right side pilot tunnel I, the left side pilot tunnel II and the middle pilot tunnel III of the arch cover part are excavated and supported, the middle temporary cross support is removed in sections, and structural arch cover lining is constructed;

[0070] S15. A left part IV-1 of a middle upper pilot tunnel of the straight wall part is excavated, a right upper pilot tunnel V of the straight wall part is further constructed and supported when the left part IV-1 of the middle upper pilot tunnel is advanced by 5 m, each cycle of the left part IV-1 of the middle upper pilot tunnel is 4.0 m, and each cycle of the right upper pilot tunnel V is 2.0 m;

[0071] S16. A right part IV-2 of the middle upper pilot tunnel of the straight wall part is excavated, a left upper pilot tunnel VI is further constructed and supported when the right part IV-2 of the middle upper pilot tunnel is advanced by 5 m, each cycle of the right part IV-2 of the middle upper pilot tunnel is 4.0 m, and each cycle of the left upper pilot tunnel VI is 2.0 m;

[0072] S17. A middle pilot tunnel VII is excavated when the middle upper pilot tunnel IV is advanced by 5 m than the middle pilot tunnel VII, and each cycle of the middle pilot tunnel VII is 4.0 m;

[0073] S18. After the middle middle pilot tunnel VII advances the right middle pilot tunnel VIII and the left middle pilot tunnel IX by 5m, excavate and support the right middle pilot tunnel VIII and the left middle pilot tunnel IX, and each cycle of the right middle pilot tunnel VIII and the left middle pilot tunnel IX is 2.0m;

[0074] S19. After the middle middle pilot tunnel VII advances the middle lower pilot tunnel X by 4-5m, excavate the middle lower pilot tunnel X, and each cycle of the middle lower pilot tunnel X is 4.0m;

[0075] S110. After the middle lower pilot tunnel X advances the right lower pilot tunnel XI and the left lower pilot tunnel XII by 5-10m, excavate and support the right lower pilot tunnel XI and the left lower pilot tunnel XII, and each cycle of the right lower pilot tunnel XI and the left lower pilot tunnel XII is 2.0m.

[0076] In particular, referring to FIG. 2A, in order to avoid damage to the upper arch cover caused by lower blasting construction, during the initial support construction of the arch cover structure 101 and the joist 102 on both sides of the arch in the steps S11 and S12, a side wall 12 is excavated 50cm thick in the upper arch cover part to the lower unexcavated part 104, so as to initially support the side wall part, the tunnel spoil (with a maximum particle size of not greater than 10cm) is layered and tamped to form a 40cm-thick spoil compaction layer, and the thickness of each layer can be 10cm, and a 10cm-thick C15 concrete layer is constructed on the top surface of the spoil compaction layer.

[0077] In particular, referring to FIG. 2B, during the construction of the right upper pilot tunnel V and the left upper pilot tunnel VI in the steps S15 and S16, the spoil compaction layer and the C15 concrete layer are cleaned in advance, and a 1.5mm-thick iron sheet layer 14 is used to cover the waterproof roll 13 below the arch cover structure, the outer edges of the iron sheet layer are wrapped with L56*36*3 angle steel 15, and during blasting, the outer side is blasted first, and then the side wall part is blasted, so as to reduce the damage to the arch cover structure and the waterproof roll caused by blasting.

[0078] S2. Referring to FIG. 4, the positions of the peripheral eye drill holes of each block are determined.

[0079] The step S2 includes the following steps:

[0080] S21. It is necessary to determine whether to use smooth blasting or pre-splitting blasting. Generally speaking, smooth blasting is used for geological conditions with high weathering degree and broken surrounding rock, and pre-splitting blasting is used for geological conditions with hard and complete surrounding rock, and preferably, the present scheme uses smooth blasting.

[0081] S22. Commonly, the interval of peripheral holes is 40-50 cm for soft rock, 45-55 cm for medium hard rock, and 50-60 cm for hard rock. The rock condition of the present example is medium hard rock, and the interval of peripheral holes is generally 45-55 cm. The interval of peripheral holes of the present example is 50 cm. In particular, to ensure the effect of smooth blasting and reduce the disturbance to the surrounding rock, uncoupling charging is needed. In addition to the longitudinal uncoupling of air interval charging in the blast hole, the transverse uncoupling coefficient requirement that the diameter of the cartridge is less than the diameter of the blast hole should also be met. The transverse uncoupling coefficient Kh = blast hole diameter / cartridge diameter, and kh is generally 1.3-2.0. The cartridge diameter of the present example is 32 mm, the borehole diameter is 42 mm, and the transverse uncoupling coefficient is 1.31, which meets the requirement.

[0082] S23. Referring to FIG. 4, to reduce the disturbance to the surrounding rock, a hole can be drilled between the peripheral holes as a damping hole.

[0083] S24. In particular, because the arch part is excavated by the double sidewall pilot hole method, to avoid damage to the temporary arch supports on both sides, two-time blasting is performed, and 70 cm of rock body close to both sides of the steel frame is reserved for secondary excavation. During construction, holes are drilled in the rock body close to the middle temporary support according to the design interval of peripheral holes, but no charging is performed. After the first blasting is completed, the 70 cm of rock body is basically loosened, and mechanical excavation can be performed. If the rock body is not loosened, secondary charging can be performed for supplementary blasting, and then mechanical excavation is performed.

[0084] S3. Determine the positions of the cut hole of each block of the arch part.

[0085] The step S3 includes the following steps:

[0086] S31. The most commonly used wedge cut blasting in throw blasting is adopted, and the cut holes are arranged in pairs in the vertical direction, and a wedge-shaped notch will be blasted after blasting.

[0087] S32. The distance h of the cut hole from the floor is preferably about 0.5-0.8 m. Too large distance will cause the failure of floor lifting. In the present example, the distance of the cut hole from the floor in the I and II zones is 0.62 m, and the distance in the III-1 zone is 0.7 m.

[0088] S33. The cut hole is about 0.2-0.3 m deeper than the auxiliary hole. The cut hole parameters depend on the surrounding rock grade, and the design parameters are shown in Table 1. In Table 1, a is the vertical distance between the cut holes, B is the horizontal distance between the cut holes, and γ is the angle between the direction of the blast hole and the working face.

[0089] Table 1. Cut hole parameters

[0090] Referring to FIG. 3A and FIG. 3B, in the embodiment, the horizontal distance B of the slotting eye is 1.5 m, the vertical distance a of the slotting eye is 0.5 m, and the angle γ between the direction of the borehole on the plane and the working face is 54°.

[0091] S4. Determine the auxiliary eye drilling position of each sub-block.

[0092] The step S4 includes the following steps:

[0093] S41. The auxiliary hole spacing is 50-90 cm, the control resistance line is 1.05-1.25 times the peripheral eye hole spacing, the large value is taken for hard rock, and the small value is taken for soft rock. The inner circle eye position is determined first according to the control resistance line, and then the remaining auxiliary eyes are arranged.

[0094] S42. The overall principle of the borehole arrangement is to minimize the number of boreholes, and the arrangement of the auxiliary eyes should be performed after the slotting eyes and the peripheral eyes are arranged. The number of drilling holes per unit area should be between 1.5-4.5 / m2. Specifically, the calculation can be performed according to the formula .

[0095] Wherein N is the number of boreholes, q is the explosive unit consumption, S is the area of the auxiliary eye excavation range, r is the linear charge density of the explosive (kg / m), and n is the borehole charging coefficient. q can be taken according to Table 2 below, and n can be taken according to Table 3 below. The value of r is determined in step S71, and it is recommended to first determine the value of N according to the approximate number of drilling holes per unit area in actual operation, and then calculate the linear charge density of the explosive through the preset N value. After iteration, the optimal economic benefit of the number of boreholes and the linear charge density is obtained.

[0096] Table 2: Explosive unit consumption value table under different rock conditions

[0097] Table 3: Borehole charging coefficient table under different rock conditions

[0098] In the embodiment, the auxiliary hole spacing is 75 cm, the control resistance line is 65 cm from the peripheral eye, the inner circle eye position is determined first according to the control resistance line, and then the remaining auxiliary eyes are arranged.

[0099] S5. Determine the bottom hole drilling position of each sub-block.

[0100] In S5, the bottom hole position needs to be determined according to the rock area that needs to be blasted on the section. In the embodiment, the bottom holes are uniformly distributed, and the spacing is between 50-65 cm.

[0101] S6. According to the confirmed drilling position, the peripheral eye, the auxiliary eye, the slotting eye, and the bottom hole are drilled.

[0102] The step S6 includes the following steps:

[0103] S61. The auxiliary eye hole depth is determined according to the cycle footage, and each cycle footage is one lattice steel frame, and the bottom over-drilling length of the hole is 0.2m-0.3m. Taking the arch part as an example, the cycle footage of the embodiment is 0.5m, and the auxiliary eye hole depth is 0.7m.

[0104] S62. The hole depth of the peripheral eye and the floor eye is 0-0.2m deeper than that of the auxiliary eye, and the hole depth of the cut eye is 0.1-0.3m deeper than that of the auxiliary eye. Taking the arch part as an example, the hole depth of the peripheral eye and the floor eye of the embodiment is 0.7m, and the hole depth of the cut eye is 1.03m, and the depth in the normal direction of the working face is 0.85m.

[0105] S63. Drilling is performed according to the positions of the peripheral eye, the auxiliary eye, the cut eye and the floor eye determined in S1-S5, and the hole depths of the peripheral eye, the auxiliary eye, the cut eye and the floor eye determined in S61 and S62. The drilling should be strictly performed according to the positions, directions and angles of the blasting design, and the drilling speed should be slow at first and then fast. During the drilling process, careful operation must be performed to prevent drill jamming, over-drilling, drill leakage and drill error.

[0106] S7. After the drilling is completed, charging is performed according to the scheme, and detonation is sequentially performed according to the excavation scheme and the positions of each detonator section.

[0107] The step S7 comprises the following steps:

[0108] S71. Calculating the single-hole charge amount and charging, comprising the following sub-steps:

[0109] The single-hole charge amount of different holes is determined, and the calculation method is referred to step S42, and the calculation formula is wherein N is the number of holes, q is the unit consumption of explosive, S is the area of the auxiliary eye excavation range, r is the linear charge density of the explosive (kg / m), n is the hole charge coefficient, and L is the hole length. Considering that the actual cartridge weight is a fixed value, only an integer number of cartridges can be loaded during charging, and it is recommended that the number of cartridges after rounding off and the linear charge density be iteratively corrected to the number of holes in step S42 during actual operation, so as to achieve optimal economic benefits; wherein the single-hole charge amount of the cut eye in the embodiment is 0.6kg, the single-hole charge amount of the auxiliary eye is 0.2kg, the single-hole charge amount of the peripheral eye is 0.2kg, and the single-hole charge amount of the floor eye is 0.2kg.

[0110] Before charging, it is necessary to check whether the hole position, depth and inclination angle meet the design requirements, whether there is blockage in the hole, whether there are falling blocks on the hole wall, and whether there is water accumulation in the hole. If it is found that the hole position and depth do not meet the design requirements, timely processing is performed, and the hole is supplemented or penetrated, and it is strictly forbidden to drill less holes and charge more.

[0111] The charge should be strictly according to the design of explosive variety, specification and quantity, not less than the charge, supercharge, and affect the blasting effect. And according to the design of the initiation device. Light blast hole air column interval charge, main hole with bulk explosive concentrated in the bottom.

[0112] Hole packing. The packing material is drilling stone powder or clay. The packing should be carried out by layering and tamping, and there should be no gap or discontinuity.

[0113] S72. Single segment maximum charge calculation. According to the formula of 13.2.4 section of "Blasting Safety Regulations" (GB6722-2014) Carry out checking calculation.

[0114] In the formula: K, a is the coefficient and attenuation coefficient related to the terrain and geology between the blasting point and the blasting object. Generally speaking, K is 200, and a is 1.8; Q is the maximum charge of single segment, unit kg; R is the distance from the blasting center point to the protected object, unit m; V is the calculated blasting vibration speed, cm / s.

[0115] According to the allowed vibration speed of the protected building, the maximum charge of single segment can be calculated, and according to the charge of single hole in S71, the number of blast holes of single segment detonator can be determined.

[0116] Referring to Figure 5, the minimum distance between the blasting center point and the civil air defense engineering to be protected is 16.9m, the maximum charge of single segment in Ⅲ-1 area is x kg, and the vibration speed control value is 1.5cm / s. Then

[0117] x = 1.397kg

[0118] x = 1.397kg

[0119] Therefore, for Ⅲ-1 area, the single hole charge of slotting eye is 0.6kg, the single hole charge of auxiliary eye is 0.2kg, the single hole charge of peripheral eye is 0.2kg, and the single hole charge of bottom plate eye is 0.2kg. Therefore, the single segment initiation of slotting eye should not exceed 2 holes, and the single segment initiation of auxiliary eye, peripheral eye and bottom plate eye should not exceed 7 holes.

[0120] S73. Delayed blasting interference vibration reduction technology is adopted to reduce the maximum vibration speed and the maximum initiation charge. The delay time interval of the previous section is relatively short. In order to avoid the influence of vibration superposition, 1, 3, 5 and 7 segments are skipped. The specific delay time is shown in the following table 4.

[0121] Table 4. Delay time table of non-electric plastic detonator

[0122] S74. Initiation of the network. The initiation of the network is in the form of "grabbing" initiation, i.e. one delay detonating tube detonator is placed in each blast hole, and then the detonating tubes are connected together, and one detonator is used to initiate the network. The detonating tube network should be connected strictly according to the design requirements, and there should be no dead knots in the detonating tube network, and there should be no joints in the blast hole, and there should be sufficient distance between the adjacent detonating tubes outside the hole. The detonating tubes are bundled into bundles according to the specified number, after the completion of the bundle, each bundle is pulled up, and it is checked whether there are unconnected (or too loose) detonating tubes on the working face. After checking that the blasting network is correct, the initiation can be carried out.

[0123] In Table 5, the type of the segmental blast hole in each section of Fig. 4 is described.

[0124] Table 5: Type and number of segmental blast holes in each blasting area

[0125] It is apparent that the foregoing description and the examples are merely illustrative and not restrictive of the disclosure, application, or uses of the present application. Although the present application has been described in detail in the examples and with reference to the accompanying drawings, the present application is not limited to the particular examples described in detail in the examples and illustrated in the accompanying drawings, but the scope of the present application will include any embodiments falling within the foregoing description and the appended claims as the best mode of carrying out the teachings of the present application presently considered.

Claims

1. A rock stratum arch cover method station blasting control construction method, comprising the following steps: S1. According to the construction blueprint, determine the excavation scheme, divide the entire station section into arch cover parts and straight wall parts, wherein the arch cover parts include three parts, excavated by double side wall pilot holes, the straight wall parts include nine parts, excavated by benching, thereby a total of 12 parts are excavated; S2. Determine the peripheral hole drilling position of each block; S3. Determine the slotting hole drilling position of each block of the arch part; S4. Determine the auxiliary hole drilling position of each block; S5. Determine the floor hole drilling position of each block, which needs to be determined according to the rock area to be blasted on the section; S6. According to the confirmed drilling position, drill the peripheral hole, auxiliary hole, slotting hole and floor hole; S7. After drilling, charge according to the scheme, and sequentially detonate according to the excavation scheme and each detonator segment position.

2. The rock strata benching method of blast control construction of a station according to claim 1, characterized in that: The step S1 specifically comprises the following sub-steps: S11. Construct the upper advanced small guide pipe of the right side pilot hole of the arch cover part, excavate the right side pilot hole I of the arch cover part, and construct the initial support of the right side pilot hole of the arch cover part; S12. Construct the upper advanced small guide pipe of the left side pilot hole of the arch cover part, excavate the left side pilot hole of the arch cover part when the right side pilot hole of the arch cover part is advanced by not less than 15m than the left side pilot hole, and construct the initial support of the left side pilot hole of the arch cover part; S13. Construct the upper advanced small guide pipe of the middle pilot hole of the arch cover part, excavate the middle pilot hole of the arch cover part by benching when the left side pilot hole of the arch cover part is advanced by not less than 15m than the middle pilot hole of the arch cover part; S14. After the excavation and support of the right side pilot hole, the left side pilot hole and the middle pilot hole of the arch cover part are completed, segmentally remove the intermediate temporary cross brace, and construct the structural arch cover lining; S15. Excavate the left part of the middle upper pilot hole of the straight wall part, and then construct the right upper pilot hole of the straight wall part and support when the left part of the middle upper pilot hole is advanced by 5m; S16. Excavate the right part of the middle upper pilot hole of the straight wall part, and then construct the left upper pilot hole and support when the right part of the middle upper pilot hole is advanced by 5m; S17. Excavate the middle middle pilot hole when the middle upper pilot hole is advanced by 4-5m; S18. Excavate the right middle pilot hole and the left middle pilot hole and support when the middle middle pilot hole is advanced by 5m than the right middle pilot hole and the left middle pilot hole; S19. Excavate the middle lower pilot hole when the middle middle pilot hole is advanced by 4-5m than the middle lower pilot hole, and the middle lower pilot hole is 4.0m per cycle footage; S110. Excavate the right lower pilot hole and the left lower pilot hole and support when the middle lower pilot hole is advanced by 5-10m than the right lower pilot hole and the left lower pilot hole.

3. The rock strata benching method of blast control construction of a station according to claim 2, characterized in that: In the initial support construction of the two side pilot holes in the steps S11 and S12, 50cm thick side walls are additionally excavated from the upper arch cover part to the lower unexcavated part to pre-advance the initial support of the side wall part, and a multi-layered compacted spoil layer is formed by using tunnel spoil layering and ramming compaction under the side wall, and a C15 concrete layer is constructed on the top surface of the compacted spoil layer.

4. The rock strata benching method of blast control construction of a station according to claim 3, characterized in that: In the step S15, S16, the right upper guide hole and the left upper guide hole are constructed, the spoil compacted layer and the C15 concrete layer are cleaned in advance, and the iron layer is used to cover the waterproof roll material, and the outer corners of the iron layer are wrapped with angle steel.

5. The rock strata benching method of blast control construction of a station according to claim 1, characterized in that: The step S2 comprises the following steps: S21. It is determined whether to use smooth blasting scheme or pre-splitting blasting, the smooth blasting is used for the geological condition of high weathering degree and broken surrounding rock, and the pre-splitting blasting is used for the geological condition of hard and complete surrounding rock; S22. The peripheral hole spacing is generally 40-50 cm for soft rock, 45-55 cm for medium-hard rock, and 50-60 cm for hard rock; S23. In order to reduce the disturbance to the surrounding rock, a hole is drilled between the peripheral hole as a damping hole; S24. Because the arch part is excavated by using the double side wall guide hole method, in order to avoid the damage to the temporary arch frame on both sides, the blasting is divided into two times, the rock body of 70 cm close to the steel frame on both sides is reserved for secondary excavation, and the hole is drilled according to the peripheral hole design spacing in the rock body close to the middle temporary support during construction, but no charge is made, after the first blasting, the 70 cm rock body is basically loose, and the mechanical excavation can be used, if there is no loosening, the second charge is made for supplement blasting, and then the mechanical excavation is used.

6. The rock strata benching method of blast control construction of a station according to claim 1, characterized in that: The step S3 comprises the following steps: S31. The wedge-shaped cutting blasting in the throwing blasting is used, the wedge-shaped cutting hole is arranged in the vertical direction, and the wedge-shaped slot is blasted after the blasting; S32. The distance between the cutting hole and the bottom plate is 0.5-0.8 m; S33. The cutting hole is 0.2-0.3 m deeper than the auxiliary hole, and the design parameters are shown in Table 1. Table 1. Trimming parameters Wherein a is the vertical distance of the cutting hole, B is the horizontal distance of the cutting hole, and γ is the angle between the hole direction on the plane and the working face.

7. The rock strata benching method of blast control construction of a station according to claim 1, characterized in that: The step S4 comprises the following steps: S41. The auxiliary hole spacing is 50-90 cm, and the control resistance line is 1.05-1.25 times of the peripheral hole spacing, and the large value is taken for hard rock and the small value is taken for soft rock; S42. The overall principle of the blast hole arrangement is to minimize the number of blast holes. The arrangement of auxiliary holes should be carried out after the arrangement of cut holes and peripheral holes. The number of holes per unit area should be between 1.5 and 4.5 per m2, and the specific value should be calculated according to the formula The calculation is performed; wherein N is the number of holes, q is the explosive unit consumption, S is the auxiliary hole excavation area, r is the linear charge density of the explosive (kg / m), and n is the hole charge coefficient.

8. The rock strata benching method of blast control construction of a station according to claim 1, characterized in that: The step S6 comprises the following steps: S61. The auxiliary hole depth is determined according to the cycle footage, and each cycle footage is a frame of grid steel frame, and the hole bottom is 0.2-0.3 m longer than the drilling length; S62. The peripheral hole and the bottom hole are 0-0.2 m deeper than the auxiliary hole, and the cutting hole bottom is 0.1-0.3 m deeper than the auxiliary hole bottom; S63. The positions of the peripheral hole, the auxiliary hole, the cutting hole and the bottom hole are determined according to S1-S5, and the peripheral hole, the auxiliary hole, the cutting hole and the bottom hole depths are determined according to S61 and S62, and the drilling is performed.

9. The rock strata bench blasting control construction method as claimed in claim 1, wherein: The step S7 comprises the following steps: S71. Calculate the single-hole charge and charge, determine the single-hole charge of different holes, the formula is Wherein N is the number of holes, q is the explosive unit consumption, S is the auxiliary hole excavation area, r is the linear charge density of the explosive (kg / m), n is the hole charge coefficient, and L is the hole length. S72. Single-stage maximum charge calculation, according to the formula in Section 13.2.4 of the "Blasting Safety Regulations" (GB6722-2014) Check, in which: K, a is the coefficient and attenuation coefficient related to the terrain, geology between the blasting point to the blasting object, generally speaking, K takes 200, a takes 1.8; Q-- single segment maximum charge, unit kg; R-- the distance from the blasting center to the protected object, unit m; V-- the calculated blasting vibration velocity, cm / s; S73. The delay blasting interference damping technology is adopted to reduce the maximum blasting vibration speed and the maximum detonating explosive amount; S74. The detonation network is laid, and the detonation is carried out after the blasting network is checked to be correct.

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