A small clear distance tunnel cycle blasting surrounding rock damage active regulation method

Abstract

The application provides a small-clearance tunnel cyclic blasting surrounding rock damage active regulation method. Through grouting filling rock mass fissure to block groundwater seepage channel, according to the position of the blast hole, the continuous and interval charging structure is respectively arranged and the initiation sequence is matched to make the blasting energy directional transmission, the arch shoulder, arch foot and middle rock area are pre-reinforced, the single section charging amount is controlled during blasting excavation and the support is applied in time. The method progressively intervenes from four dimensions of water-induced conditions, energy source, structural weakness and expansion path of damage, forms a complete active regulation closed loop, and significantly reduces the cumulative damage degree of surrounding rock.

Description

Technical Field

[0001] This invention relates to the field of tunnel engineering technology, and in particular to a method for actively controlling the damage of surrounding rock in cyclic blasting of tunnels with small clearance. Background Technology

[0002] Small-clearance tunnels are a common tunnel structure in mountainous highway and railway construction. Due to the limitations of terrain and geological conditions, the clearance between the two tunnels is much smaller than that of conventional separated tunnels. Drill-and-blast method remains the primary excavation method used in mountain tunnel construction due to its adaptability to various surrounding rock conditions and its economic efficiency. Tunnel construction typically requires dozens or even hundreds of cycles of blasting excavation. Each blast, while breaking the rock at the tunnel face, also propagates stress waves to the surrounding rock outside the excavation outline, disturbing the remaining surrounding rock.

[0003] During cyclic blasting excavation, the blasting disturbances from the initial and subsequent tunnels can easily create a cumulative effect in the interbedded rock area. While the damage to the surrounding rock caused by a single blast is limited, the repeated action of multiple blasts causes the micro-fractures within the rock mass to gradually expand and connect, and the degree of damage accumulates continuously with the increase in the number of excavation cycles. This cumulative damage leads to a decrease in the integrity of the surrounding rock and a gradual deterioration of its mechanical parameters, which, in severe cases, may affect the overall stability of the tunnel and construction safety. How to effectively control the cumulative damage to the surrounding rock under cyclic blasting conditions is a key technical challenge in the construction of tunnels with small clearances.

[0004] From the current state of the industry, existing methods for controlling surrounding rock damage mainly focus on two aspects: controlling blasting vibration intensity and passive support after excavation. However, a systematic technical solution has not yet been formed to address the unique characteristics of small-clearance tunnels, such as the superposition of disturbances between two tunnels, the coupling effect of groundwater and blasting stress waves, and the spatiotemporal evolution of damage accumulation. The industry still lacks effective methods to support the active control of surrounding rock damage. Summary of the Invention

[0005] In view of the above problems, a method for active control of surrounding rock damage in small-clearance tunnel cyclic blasting is proposed to overcome or at least partially solve these problems, including:

[0006] Grouting is used to fill the original fissures in the surrounding rock mass of the tunnel.

[0007] For the surrounding rock after grouting, blast holes with continuous charge structure are set at the center of the tunnel face, and blast holes with interval charge structure are set at the tunnel excavation outline. The detonation sequence is set so that the blast holes at the center of the tunnel face are detonated before the blast holes at the excavation outline.

[0008] Pre-reinforcement is carried out on the arch shoulder, arch foot, and interbedded rock areas of the surrounding rock.

[0009] After pre-reinforcement of the arch shoulder, arch foot, and interbedded rock areas, blasting excavation is carried out. During each blasting excavation, the net distance between the advance tunnel and the follow tunnel, the surrounding rock grade, and the allowable peak vibration velocity are obtained. Based on the net distance, surrounding rock grade, and allowable peak vibration velocity, the maximum charge threshold for a single section is determined. The actual charge amount for a single section is set to not exceed the maximum charge threshold for a single section. After each blasting excavation is completed, a support structure is applied to the excavation face.

[0010] Optionally, cement grout is used for grouting, with a water-cement ratio ranging from 0.6:1 to 1.2:1, a grouting pressure ranging from 0.5MPa to 2.0MPa, grouting holes arranged in a quincunx pattern, with a spacing of 1.0m to 2.0m, and the grouting reinforcement depth extending at least 3m beyond the tunnel excavation outline. The grouting is completed when the grouting pressure reaches the design final pressure and the grouting volume is less than the preset threshold.

[0011] Optionally, when the borehole at the center of the tunnel face adopts a continuous charge structure, the ratio of the charge diameter to the borehole diameter is within the range of the radially uncoupled charge coefficient. A concentrated charge pack is set at the bottom of the borehole at the center of the tunnel face. The concentrated charge pack is filled with a predetermined proportion of the designed charge within a predetermined length range at the bottom of the borehole. When the borehole at the excavation outline adopts an intermittent charge structure, the intermittent material is air or an inert solid medium, and the length of a single intermittent section is determined according to the borehole depth and the number of charge sections.

[0012] Optionally, the detonation sequence is set as follows: the slotted hole at the center of the working face is detonated first to create a free face. After the slotted hole is detonated, the auxiliary holes, peripheral holes and bottom edge holes are detonated in sequence, and the delay time between adjacent sections meets the conditions for staggered propagation of blasting vibration waves.

[0013] Optionally, pre-reinforcement of the surrounding rock at the arch shoulder, arch foot, and interbedded rock area includes: installing advanced small guide pipes and grouting at the arch shoulder and arch foot, with the outer insertion angle of the advanced small guide pipes determined according to the tunnel excavation outline and the circumferential spacing determined according to the degree of fracture of the surrounding rock; and installing tie rods in the interbedded rock area, with the tie rods penetrating the interbedded rock and applying prestress.

[0014] Optionally, determining the maximum charge threshold for a single segment includes: during the initial blasting excavation, selecting a site coefficient and attenuation index from empirical values ​​of blasting vibration attenuation parameters for the same level of surrounding rock as specified in the blasting safety regulations, based on the surrounding rock level; calculating the maximum charge threshold for a single segment for the initial blasting based on the site coefficient, attenuation index, allowable peak vibration velocity, and distance from the blast source to the monitoring point; and during subsequent cyclic blasting excavation, fitting a corrected site coefficient and attenuation index based on the on-site blasting vibration monitoring data of each completed blast, and updating the maximum charge threshold for a single segment based on the corrected site coefficient and attenuation index.

[0015] Optionally, the application of the support structure includes: spraying a first layer of concrete onto the surface of the excavation face; erecting a steel arch frame along the tunnel excavation outline on the surface of the first concrete layer; drilling anchor bolt holes into the surrounding rock at the arch crown, arch shoulder, arch waist, and arch foot positions of the excavation face and installing system anchor bolts; spraying a second layer of concrete to the designed thickness after the steel arch frame and system anchor bolts are installed; after the second concrete layer is sprayed, drilling anchor bolt holes into the surrounding rock at the arch foot position and installing anchor bolts, wherein the installation direction of the anchor bolts is at an inclined angle to the horizontal plane, and the number of anchor bolts installed at the same arch foot position is determined according to the degree of fracture of the surrounding rock at the arch foot.

[0016] Optionally, before grouting and filling the original fissures in the surrounding rock mass of the tunnel, the following steps are also included: sampling the surrounding rock of the tunnel and conducting SHPB dynamic impact tests to obtain the dynamic stress-strain curves and failure modes of the surrounding rock under different impact velocities, establishing a numerical model consistent with the size of the SHPB test device, and adjusting the parameters of the RHT dynamic constitutive model with the dynamic stress-strain curves and failure modes as calibration targets.

[0017] Optionally, after adjusting the parameters of the RHT dynamic constitutive model, the following steps are also included: establishing a numerical model for cyclic blasting excavation of a small-clearance tunnel based on the preliminary borehole design scheme; substituting the calibrated RHT dynamic constitutive model parameters into the numerical model; simulating multiple blasting excavation processes of the lead tunnel and the follow tunnel; extracting damage variable cloud maps and damage depth data; identifying locations where the damage variable exceeds a preset damage threshold and the damage depth exceeds a preset depth threshold as damage-sensitive areas requiring pre-reinforcement; and adjusting the charge structure and detonation sequence of the boreholes at the center of the tunnel face and the boreholes at the excavation outline based on the simulation results.

[0018] Optionally, the method also includes: obtaining the curves of the deformation of the surrounding rock and the stress on the support structure with the excavation cycle through numerical simulation of cyclic blasting, determining the construction parameters of the support structure based on the curves, and determining the specifications of the support structure so that the bearing capacity of the support structure is greater than the simulated surrounding rock pressure.

[0019] This invention achieves proactive control of surrounding rock damage in cyclic blasting construction of tunnels with small clearances. Grouting pretreatment fills the original fissures in the rock mass and blocks groundwater seepage channels, fundamentally eliminating the promoting effect of hydrodynamic pressure caused by blasting stress waves on fissure propagation. Based on the difference in energy transmission paths between blast holes at the center of the tunnel face and those at the excavation outline, continuous and intermittent charging structures are respectively set up and coordinated with a center-first detonation sequence, ensuring blasting energy is directionally transmitted towards the rock mass to be excavated, reducing energy dissipation towards the remaining surrounding rock. Pre-reinforcement is applied at the arch shoulder, arch foot, and interbedded rock areas to enhance the initial damage resistance of damage-sensitive areas. For each blasting excavation, the maximum charge threshold for a single section is determined based on the clearance between the preceding and following tunnels, the surrounding rock grade, and the allowable peak vibration velocity, and the actual charge is controlled to not exceed this threshold. Simultaneously, support structures are promptly applied after each blasting excavation to inhibit further propagation of surrounding rock fissures. The above four steps run through the entire construction process, and progressive intervention is carried out from four dimensions: water-induced conditions, energy sources, structural weaknesses and propagation paths. This forms a complete active control closed loop for the damage to the surrounding rock caused by cyclic blasting in tunnels with small clearance, which significantly reduces the cumulative damage to the surrounding rock caused by cyclic blasting. Attached Figure Description

[0020] To more clearly illustrate the technical solution of the present invention, the accompanying drawings used in the description of the present invention 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.

[0021] Figure 1 This is a flowchart of a method for actively controlling the damage of surrounding rock in cyclic blasting of tunnels with small clearance provided by an embodiment of the present invention;

[0022] Figure 2 This is a schematic diagram of the SHPB dynamic impact testing device provided in an embodiment of the present invention;

[0023] Figure 3 This is a graph showing the relationship between peak vibration velocity and proportional distance in the interbedded rock provided in an embodiment of the present invention;

[0024] Figure 4 This is a graph showing the variation of peak vibration velocity at various measuring points in the interbedded rock with blasting advance, provided in an embodiment of the present invention. Detailed Implementation

[0025] To make the above-mentioned objects, features, and advantages of the present invention more apparent and understandable, the present invention will be further described in detail below with reference to the accompanying drawings and specific embodiments. Obviously, the described embodiments are only some, not all, of the embodiments of the present invention. All other embodiments obtained by those skilled in the art based on the embodiments of the present invention without creative effort are within the scope of protection of the present invention.

[0026] Reference Figures 1 to 4 This invention provides a method for actively controlling the damage of surrounding rock in cyclic blasting of tunnels with small clearance, which may specifically include:

[0027] Grouting is used to fill the original fissures in the surrounding rock mass of the tunnel.

[0028] For the surrounding rock after grouting, blast holes with continuous charge structure are set at the center of the tunnel face, and blast holes with interval charge structure are set at the tunnel excavation outline. The detonation sequence is set so that the blast holes at the center of the tunnel face are detonated before the blast holes at the excavation outline.

[0029] Pre-reinforcement is carried out on the arch shoulder, arch foot, and interbedded rock areas of the surrounding rock.

[0030] After pre-reinforcement of the arch shoulder, arch foot, and interbedded rock areas, blasting excavation is carried out. During each blasting excavation, the net distance between the advance tunnel and the follow tunnel, the surrounding rock grade, and the allowable peak vibration velocity are obtained. Based on the net distance, surrounding rock grade, and allowable peak vibration velocity, the maximum charge threshold for a single section is determined. The actual charge amount for a single section is set to not exceed the maximum charge threshold for a single section. After each blasting excavation is completed, a support structure is applied to the excavation face.

[0031] Specifically, before tunnel construction, grouting is performed on the surrounding rock within the tunnel excavation area. Grouting material fills the original fissures in the rock mass, sealing the originally interconnected fissure network into isolated pores, thereby blocking the seepage channels of groundwater along the fissure surfaces within the rock mass. The grouting holes are laid outwards along the tunnel excavation outline, with the final depth extending beyond the excavation outline. The selection of grouting material is determined comprehensively based on the degree of fissure development in the surrounding rock and the groundwater occurrence conditions. Grouting is considered complete when the grouting pressure reaches the designed final pressure and the grouting volume per unit time remains below the preset termination threshold.

[0032] The mechanism by which grouting blocks groundwater seepage channels lies in eliminating the coupling effect between blasting stress waves and fissure water. In ungrouted surrounding rock, groundwater exists within rock fissures. When blasting stress waves propagate to the water-bearing fissure surface, the instantaneous high pressure caused by the stress waves acts on the fissure water, forcing it into the fissure tip and creating a water wedge effect. Simultaneously, the stress waves directly damage the rock mass framework. The water wedge effect and the stress wave damage effect mutually promote each other. Under the continuous action of the stress waves, the fissure water is repeatedly forced into the fissure tip, and the stress intensity factor at the fissure tip gradually increases. When it exceeds the fracture toughness of the rock mass, the fissure expands. The expanded fissure then provides a channel for more fissure water to enter, forming a positive feedback loop that accelerates damage.

[0033] After grouting, the grout fills the rock fissures and solidifies. The space occupied by the solidified grout in the fissures no longer has the conditions for the existence of free water. When the blasting stress wave propagates to the filled fissures, the carrier of the dynamic water pressure effect is eliminated because there is no free water medium in the fissures. The preconditions for the water wedge splitting effect no longer hold, and the positive feedback mechanism of water-induced damage acceleration is weakened from the root.

[0034] After grouting the surrounding rock, the layout of blast holes and the design of the charging structure are carried out at the tunnel face. Based on their location on the face, the blast holes are divided into two categories: blast holes at the center of the face and blast holes along the excavation outline. Blast holes at the center of the face are located in the central area of ​​the face. These blast holes are surrounded by intact rock mass, with only one free face facing the face. The rock mass exerts a significant clamping force on these blast holes. After the explosives inside the blast hole detonate, they need to overcome the clamping force of the rock mass and break up a cavity facing the face to create a new free face. Blast holes at the center of the face use a continuous charging structure. A continuous charging structure means that explosive cartridges are continuously loaded along the axial direction of the blast hole without any spacer between cartridges. During blasting, the cartridges continuously detonate along the axial direction of the blast hole, and the energy is concentrated and released radially, making the pressure of the blast on the surrounding rock mass more concentrated.

[0035] The blast holes at the excavation outline location are situated in the area surrounding the working face. One side of these blast holes is the fractured rock mass to be excavated, and the other side is the surrounding rock to be preserved. After the explosives detonate, the energy is transmitted simultaneously in both directions: the side to be excavated and the side to be preserved. If a continuous charge structure is used for the blast holes at the excavation outline location, the blasting energy is released uniformly along the radial direction of the blast hole, and the energy transmitted towards the preserved surrounding rock will cause additional damage to the preserved surrounding rock. Alternatively, if a spaced charge structure is used for the blast holes at the excavation outline location, this means that the explosive cartridges are loaded in segments along the axial direction of the blast hole, with spacers between adjacent cartridges. These spacers are not filled with explosives; the spacer medium is air or a spacer made of inert material. During blasting, the explosive cartridges detonate in segments along the axial direction of the blast hole. Due to the presence of the spacers, the distribution of blasting energy along the radial direction of the blast hole tends to be more uniform, and the blasting pressure acting on the excavation outline surface is more evenly distributed along the outline direction, thus reducing the proportion of energy transmitted towards the preserved surrounding rock.

[0036] After the explosive charge structure is determined, the detonation sequence is set so that the boreholes at the center of the face detonate before those at the excavation outline. Among the boreholes at the center of the face, the cut hole is the first to detonate. Located in the central region of the face, the cut hole creates a cavity in the rock mass facing the face; the surface of this cavity becomes the new free surface that can be utilized by subsequent borehole blasts. After the cut hole detonates, the remaining boreholes at the center of the face and those at the excavation outline detonate sequentially. The explosive energy generated by the subsequently detonated boreholes is directionally transmitted and concentratedly released towards the created free surface.

[0037] After completing the design of the borehole charge structure and setting the detonation sequence, pre-reinforcement was implemented on the arch shoulder, arch foot, and interbedded rock areas of the surrounding rock. The arch shoulder is the transition arc section between the arch and sidewall of the tunnel excavation outline, and the arch foot is the junction of the sidewall and bottom slab of the tunnel excavation outline. Both of these are geometric inflection points of the excavation outline. When the blasting stress wave propagates to the geometric inflection point, the abrupt curvature change of the wavefront at the inflection point causes diffraction and reflection of the stress wave, making the dynamic stress concentration in the surrounding rock near the inflection point higher than that in the straight outline area. Under the repeated loading of multiple blasting cycles, the damage accumulation rate and degree at the geometric inflection point are greater than those in other outline areas. The interbedded rock area is the rock column between the first and second tunnels. During construction, it is subjected to the alternating and superimposed effects of blasting disturbances from both directions. The bidirectional disturbance causes the micro-cracks inside the interbedded rock to expand alternately in both directions, making the damage accumulation path more complex and the damage evolution rate faster. Pre-reinforcement improves the initial damage resistance of the three types of damage-sensitive areas by applying reinforcement structures in advance.

[0038] After pre-reinforcement of the arch shoulder, arch foot, and interbedded rock areas, blasting excavation is carried out. Before each blasting excavation, three parameters are obtained: the current net distance between the preceding and following tunnels, the surrounding rock grade, and the allowable peak vibration velocity. The net distance is the shortest distance between the excavation outlines of the preceding and following tunnels, and this value changes gradually as construction progresses; the surrounding rock grade is determined based on the comprehensive evaluation results of rock mass integrity, uniaxial saturated compressive strength, and the degree of development of structural surfaces in the tunnel geological survey report; the allowable peak vibration velocity is the upper limit of the blasting vibration velocity control set according to the tunnel structural safety requirements and the protection needs of adjacent existing facilities.

[0039] The maximum charge threshold for a single blast segment is determined based on three parameters: net distance, surrounding rock grade, and allowable peak vibration velocity. The determination method involves using the net distance and surrounding rock grade to determine the attenuation characteristic parameters of the blast vibration wave propagating from the blast source to the monitoring point. Then, using the attenuation characteristic parameters, allowable peak vibration velocity, and the distance from the blast source to the monitoring point as inputs, the allowable maximum charge value for a single blast segment is calculated. This calculated value is the maximum charge threshold for a single blast segment. The actual charge amount for each delay segment in the blast design is compared segment by segment with the maximum charge threshold for a single segment. If the charge amount for a certain delay segment exceeds the threshold, the charge amount for that delay segment is reduced to below the threshold, or the delay segment is split into two or more delay segments, ensuring that the charge amount for each segment does not exceed the maximum charge threshold for a single segment.

[0040] After each blasting excavation, a support structure is applied to the excavation face. The support structure is constructed by building a load-bearing structural system layer by layer on the surface of the excavation face. The support structure applies radial constraint force to the excavation face, limiting the deformation of the surrounding rock in the direction of the excavation space after blasting, and inhibiting the further expansion and interconnection of micro-cracks inside the surrounding rock caused by blasting during the stress adjustment process.

[0041] In one or more embodiments of the present invention, the grouting filling uses cement grout with a water-cement ratio ranging from 0.6:1 to 1.2:1, a grouting pressure ranging from 0.5MPa to 2.0MPa, grouting holes arranged in a quincunx pattern with a spacing of 1.0m to 2.0m, and the grouting reinforcement depth extending beyond the tunnel excavation outline by no less than 3m. The grouting completion criterion is that the grouting pressure reaches the designed final pressure and the grouting volume is less than a preset threshold.

[0042] In practical applications, the specific value of the water-cement ratio can be determined through on-site grouting tests before grouting operations.

[0043] Specifically, test holes were drilled at representative locations outside the tunnel excavation outline. Cement grout with different water-cement ratios was used for test injection. The relationship curve between the grout diffusion radius and the grouting pressure under each water-cement ratio was recorded. The corresponding water-cement ratio that allows the grout diffusion radius to cover the distance between adjacent grouting holes within the grouting pressure range was selected as the construction water-cement ratio.

[0044] The grouting holes are arranged in a quincunx pattern along the outer edge of the tunnel excavation outline, with the spacing between adjacent holes ranging from 1.0m to 2.0m. The spacing is determined based on the degree of fissure development in the surrounding rock. The method involves drilling core samples before grouting, measuring the average spacing of fissures and the opening width of the fissure surface in the core samples. When the average fissure spacing is less than the core diameter and the fissure opening width is greater than the minimum width that cement grout can be injected, the larger value within this range is used for the grouting hole spacing. When the average fissure spacing is greater than the core diameter or the fissure opening width is close to the minimum width that cement grout can be injected, the smaller value within this range is used for the grouting hole spacing. The quincunx arrangement ensures that the positions of adjacent rows of grouting holes are staggered in the circumferential direction, and the grout diffusion area of ​​each grouting hole overlaps and covers the grout diffusion areas of adjacent grouting holes, reducing grouting blind spots caused by insufficient diffusion radius of a single grouting hole.

[0045] The grouting pressure ranges from 0.5 MPa to 2.0 MPa. During grouting, the pressure is applied in stages: Initially, the pressure is set at 0.5 MPa, and grout is continuously injected while monitoring the flow rate. When the average flow rate drops to half of the initial flow rate over 10 minutes, the pressure is increased by one step, with each step increasing by 0.3 MPa to 0.5 MPa. After each step increase, the flow rate is monitored again. When the flow rate drops to half of the initial flow rate again, the pressure is increased to the next step. Grouting stops when the final design pressure is reached and the flow rate drops below the preset termination threshold and remains at this level for at least 5 minutes. If the pressure reaches 2.0 MPa but the flow rate is still above the preset termination threshold, grouting continues at 2.0 MPa until the flow rate drops below the preset termination threshold.

[0046] The final depth of the grouting holes should extend at least 3 meters beyond the tunnel excavation outline. The specific value of the final depth is determined comprehensively based on the tunnel excavation height and the depth of fissure development in the surrounding rock, and the distance extending outward from the excavation outline should be no less than 1 / 3 of the tunnel excavation height. After grouting, the cement slurry is allowed to solidify in the fissures. The solidification time is determined based on the water-cement ratio of the cement slurry and the ambient temperature. During the solidification period, no blasting operations adjacent to the grouting area are carried out. After the grouting has solidified, the fissure filling condition is observed by drilling inspection holes and taking core samples. When the volume ratio of cement slurry-filled rock in the core sample meets the preset filling rate, subsequent blast hole layout and charge structure design can be carried out in the grouting area.

[0047] In one or more embodiments of the present invention, when the blast hole at the center of the tunnel face adopts a continuous charging structure, the ratio of the charge diameter to the blast hole diameter is within the range of the radially uncoupled charging coefficient. A concentrated charge pack is provided at the bottom of the blast hole at the center of the tunnel face. The concentrated charge pack is filled with a predetermined proportion of the designed charge within a predetermined length range at the bottom of the blast hole. When the blast hole at the excavation contour position adopts an intermittent charging structure, the intermittent material is air or an inert solid medium, and the length of a single intermittent segment is determined according to the blast hole depth and the number of charge segments.

[0048] The continuous charge structure can be set up as follows: the explosive cartridges are continuously loaded along the borehole axis from the borehole opening to the bottom, with adjacent cartridges connected end-to-end. The ratio of the charge diameter to the borehole diameter is taken within the range of the radial decoupling charge coefficient. The radial decoupling charge coefficient is defined as the ratio of the borehole diameter to the charge diameter. The specific value of this coefficient is determined based on the uniaxial saturated compressive strength of the surrounding rock and the explosive detonation velocity. The determination method is to input the ratio of the uniaxial saturated compressive strength of the surrounding rock to the explosive detonation velocity, and select the corresponding coefficient value from a preset radial decoupling coefficient value table. The function of the radial decoupling charge coefficient is to leave a circumferential gap between the borehole wall and the explosive cartridge. During blasting, the shock wave generated by the explosive detonation first attenuates through the air layer in the circumferential gap before acting on the borehole wall, thus reducing the peak pressure acting on the borehole wall, prolonging the pressure rise time, and making the pressure time distribution more gradual, thereby reducing the impact damage to the surrounding rock mass.

[0049] Concentrated explosive charges can be set up as follows: a predetermined proportion of the designed explosive charge in the borehole is concentrated and filled within a predetermined length range at the bottom of the borehole. This predetermined proportion is 30%-40% of the total explosive charge in the borehole, and the predetermined length is 1 / 3 of the borehole depth. During filling, the concentrated explosive charge is first filled to the bottom of the borehole, and then the remaining explosive charge is filled in the remaining section between the concentrated explosive charge and the borehole opening using a continuous charging structure. The function of the concentrated explosive charge is to concentrate the blasting energy at the bottom of the borehole, so that the blasting pressure on the rock mass at the bottom of the borehole is higher than that on the rock mass above the borehole, enhancing the fragmentation effect of the rock mass at the bottom of the borehole, which is beneficial for overcoming the clamping effect of the rock mass at the center of the working face and creating a free surface towards the working face.

[0050] The spaced-charge structure can be set up as follows: the designed charge amount of the borehole is divided into multiple charge segments along the borehole axis, and spacer segments are set between adjacent charge segments. No explosives are filled in the spacer segments. The medium filling the spacer segments is air or a spacer made of an inert solid material. Plastic or wooden spacers can be used for the inert solid spacers. The length of a single spacer segment is determined based on the borehole depth and the number of charge segments. The determination method is to divide the remaining length after subtracting the length of the borehole clogging segment and the length of the reserved segment at the bottom from the borehole depth by the sum of the number of charge segments and the number of spacer segments. The quotient obtained is the baseline value for the length of each charge segment and the length of each spacer segment. The charge length of each segment and the spacer length of each segment are both taken from this baseline value.

[0051] The intermittent charge structure distributes the explosive in segments along the axial direction of the borehole. During blasting, each charge segment detonates separately. The distribution of blasting energy along the axial direction of the borehole tends to be uniform due to the presence of the intermittent segments. The blasting pressure acting on the excavation profile is more evenly distributed along the profile direction, reducing the damage to the remaining surrounding rock caused by local blasting pressure concentration. At the same time, the proportion of energy transferred to the remaining surrounding rock is lower than that of the continuous charge structure.

[0052] In one or more embodiments of the present invention, the detonation sequence is set as follows: the slotting hole at the center of the working face is detonated first to create a free surface. After the slotting hole is detonated, the auxiliary holes, peripheral holes and bottom edge holes are detonated in sequence. The delay time between adjacent sections satisfies the condition for staggered propagation of blasting vibration waves.

[0053] The detonation sequence is set according to the principle that the blast holes at the center of the tunnel face are detonated before those at the excavation outline. Among the blast holes at the center of the tunnel face, the cut hole is detonated first. Located in the central region of the tunnel face, the cut hole creates a cavity facing the tunnel face within the rock mass after detonation. The surface of this cavity becomes a new free surface that can be utilized for subsequent blasting. The blasting stress wave is reflected when it reaches the free surface. The compression wave reflected from the free surface is unloaded and transformed into a tensile wave that propagates into the rock mass. The tensile wave superimposes with the subsequently transmitted incident compression wave, forming a tensile stress zone near the free surface. When the tensile stress exceeds the tensile strength of the rock, the rock undergoes tensile failure in this area. The existence of the free surface thus enhances the rock-breaking effect of the blast.

[0054] After the cut-out holes are detonated, the auxiliary holes at the center of the tunnel face are detonated. These auxiliary holes are located between the cut-out holes and the blast holes at the excavation outline. Detonation occurs under the condition that the free surface created by the cut-out holes has already formed. The blasting energy from the auxiliary holes is released directionally towards the free surface created by the cut-out holes, further expanding the range of the free surface. After the auxiliary holes are detonated, the peripheral holes at the excavation outline are detonated. These peripheral holes are adjacent to the tunnel excavation design outline and detonate under the condition of the expanded free surface. The blasting energy is released evenly along the excavation outline, completing the shaping of the tunnel cross-section outline. After the peripheral holes are detonated, the bottom edge holes are detonated. These bottom edge holes are located in the bottom area of ​​the tunnel face and detonate under the condition that the cross-section outline has been basically formed, completing the fracturing of the rock mass at the bottom of the tunnel face and forming the tunnel floor outline.

[0055] The delay time between adjacent segments is set based on the staggered propagation condition of blasting vibration waves. The staggered propagation condition refers to the condition where the peak values ​​of seismic waves generated by blasting in each segment do not overlap on the time axis when they propagate to the same monitoring location in the surrounding rock. The delay time is determined as follows: the distance from each blast hole to the monitoring point in the surrounding rock is obtained; the arrival time of the blasting seismic waves from the blast source to the monitoring point is calculated based on the propagation velocity of seismic waves in the surrounding rock; and the difference in arrival times between adjacent segments is taken as the minimum delay time, ensuring that the peak value of the seismic wave from the first blasted segment passes the monitoring point before the peak value of the seismic wave from the subsequent blasted segment arrives at the same monitoring point. The delay time that satisfies the staggered propagation condition of blasting vibration waves allows the blasting vibrations from each segment to be released in a dispersed manner over time, avoiding the simultaneous arrival of multiple seismic wave peaks at the same location in the surrounding rock, which would cause vibration superposition and amplification, thereby reducing the disturbance intensity of blasting vibrations to the surrounding rock and adjacent structures.

[0056] In one or more embodiments of the present invention, pre-reinforcement of the arch shoulder, arch foot, and interbedded rock area of ​​the surrounding rock includes: installing advanced small guide pipes and grouting at the arch shoulder and arch foot, wherein the outer insertion angle of the advanced small guide pipes is determined according to the tunnel excavation outline and the circumferential spacing is determined according to the degree of fracture of the surrounding rock; and installing tie rods in the interbedded rock area, wherein the tie rods penetrate the interbedded rock and apply prestress.

[0057] Pre-reinforcement measures can include two categories: the first category is to install advanced small guide pipes and grout at the arch shoulder and arch foot, and the second category is to install tie rods and apply prestress in the interbedded rock area.

[0058] Specifically, the installation of advance guide pipes at the arch shoulder and arch foot can be carried out as follows: Before the tunnel face is excavated to the pre-reinforced section, hollow steel pipes are drilled into the surrounding rock at the arch shoulder and arch foot along the outer side of the tunnel excavation outline. The drilling direction of the hollow steel pipe forms an inclined angle with the tunnel excavation axis, and this inclined angle is the outer insertion angle of the advance guide pipe. The outer insertion angle is determined based on the outer expansion range of the tunnel design excavation outline and the extension length of the advance guide pipe in the surrounding rock, so that the final hole position of the advance steel pipe falls in the surrounding rock outside the tunnel excavation outline. The spacing of the pre-grouted small guide pipes along the tunnel circumference is determined based on the degree of rock fragmentation at the arch shoulder and arch foot. The determination method is to drill core samples from the surrounding rock at the arch shoulder and arch foot before pre-reinforcement construction, and measure the average spacing of the cracks in the core samples. When the average spacing of the cracks is less than the preset spacing benchmark value, the circumferential spacing is taken as a smaller value, so that the grout solidified bodies formed after grouting of two adjacent pre-grouted small guide pipes overlap each other in the circumferential direction to form a continuous arch-shaped reinforcement zone. When the average spacing of the cracks is greater than the preset spacing benchmark value, the circumferential spacing is taken as a larger value, and the grout solidified bodies formed after grouting of two adjacent pre-grouted small guide pipes each independently undertake the reinforcement function of that part.

[0059] After the pre-drilled and installed guide pipes are installed, cement grout is injected into the surrounding rock fissures at the arch shoulder and arch foot through the guide pipes. The grouting pressure is set to ensure that the grout spreads along the fissure surface and fills the fissures in the surrounding rock around the guide pipes. The grouting is considered complete when the grouting pressure reaches the set value and the grout volume per unit time is lower than the preset termination threshold. After the grout solidifies in the fissures, a solidified grout body is formed in the surrounding rock at the arch shoulder and arch foot, spreading along the direction of the fissure surface with each guide pipe as the center. This solidified grout body cements the rock blocks that were originally separated by fissures into a whole, improving the integrity and damage resistance of the surrounding rock at the arch shoulder and arch foot.

[0060] The installation of tie rods in the interbedded rock area can be carried out as follows: After both the initial and subsequent tunnels have been excavated through the pre-reinforced sections, a through hole is drilled from one excavation face of the interbedded rock to the other excavation face, penetrating the interbedded rock column. The drilling direction of the through hole is perpendicular to the extension direction of the interbedded rock column. After the through hole is completed, the anchor rod is inserted into the through hole, with both ends of the anchor rod extending out of the excavation faces of the initial and subsequent tunnels, respectively. A pad and fastener are then installed sequentially at both ends of the anchor rod. By applying torque to the fasteners, the anchor rod is subjected to tension. The tension of the anchor rod is transmitted to the excavation faces of the initial and subsequent tunnels through the pads, forming a pre-stress along the anchor rod axis inside the interbedded rock column. The pre-compression stress puts the interbedded rock column under pressure before it is disturbed by the blasting of the preceding and following tunnels. When the tensile stress wave generated by the blasting disturbance propagates to the interbedded rock column, the tensile stress wave needs to overcome the pre-compression stress before the internal cracks of the rock column can open and expand. The pre-compression stress thus improves the ability of the interbedded rock column to resist bidirectional blasting disturbance.

[0061] In one or more embodiments of the present invention, determining the single-segment maximum charge threshold includes: during the initial blasting excavation, selecting a site coefficient and attenuation index from empirical values ​​of blasting vibration attenuation parameters for the same level of surrounding rock as specified in the blasting safety regulations, based on the surrounding rock level; calculating the single-segment maximum charge threshold for the initial blast based on the site coefficient, attenuation index, allowable peak vibration velocity, and distance from the blast source to the monitoring point; during subsequent cyclic blasting excavation, fitting a corrected site coefficient and attenuation index based on the on-site blasting vibration monitoring data of each completed blast, and updating the single-segment maximum charge threshold based on the corrected site coefficient and attenuation index.

[0062] In determining the maximum charge threshold for a single section before each blasting excavation, different parameter sources are used to calculate the threshold depending on whether the blast is the first blast or a subsequent cycle blast.

[0063] The empirical relationship between peak blast vibration velocity and charge amount is as follows:

[0064]

[0065]

[0066] In the formula, PPV is the peak vibration velocity at the monitoring point, K is the site coefficient, Q is the charge amount per segment, R is the distance from the blast source to the monitoring point, and α is the attenuation index.

[0067] To solve for the threshold of maximum charge in a single segment, the above relationship is transformed algebraically, and the charge Q in a single segment is expressed as a dependent variable determined by the peak vibration velocity (PPV), the site coefficient (K), the attenuation index (α), and the distance R from the blast source to the monitoring point.

[0068] The specific transformation process is as follows: First, take the logarithm of both sides of the equation to obtain a linear relationship between the logarithm of peak vibration velocity and the logarithm of proportional distance; subtract the logarithm of the site coefficient from the logarithm of the peak vibration velocity, and divide the difference by the attenuation exponent α to obtain the logarithm of the proportional distance; subtract the logarithm of the proportional distance from the logarithm of the distance R from the blast source to the monitoring point, and multiply the difference by two to obtain the logarithm of the single-segment charge Q; then, perform an exponential operation on this logarithm to obtain the calculated value of the single-segment charge Q. Substitute the allowable peak vibration velocity as the PPV value into the above algebraically transformed equation, and the calculated Q value is the threshold for the maximum single-segment charge.

[0069] For the initial blasting excavation, since no on-site blasting vibration monitoring data for this tunnel project has been accumulated, the site coefficient K and attenuation index α are selected from the empirical values ​​of blasting vibration attenuation parameters for the same level of surrounding rock as specified in the blasting safety regulations. The blasting safety regulations list an empirical value table of blasting vibration attenuation parameters corresponding to different surrounding rock levels. This table is derived through statistical regression analysis of a large amount of measured blasting vibration data from similar projects. The selection method is as follows: using the tunnel surrounding rock level as an index, the suggested values ​​of the site coefficient and attenuation index corresponding to that rock level are retrieved from the empirical value table. The retrieved suggested value of the site coefficient is used as the site coefficient K for the initial blast, and the retrieved suggested value of the attenuation index is used as the attenuation index α for the initial blast. These parameters, along with the allowable peak velocity (PPV) and the distance R from the blast source to the monitoring point, are substituted into the algebraically transformed formula to calculate the maximum charge threshold for a single section of the initial blast.

[0070] For subsequent blasting excavations following the initial blast, on-site blasting vibration monitoring data from previous blasts has been accumulated. This data includes the actual charge amount at each segment of each blast, the distance from the blast source to the monitoring point, and the measured peak velocity (PPV) at the monitoring point. This accumulated monitoring data is used to fit and correct the site coefficient K and the attenuation index α. The correction method is as follows: the actual charge amount Q and the distance R from the blast source to the monitoring point at each segment of each blast are converted into proportional distances (R divided by the square root of Q). The proportional distances and the corresponding measured PPVs are plotted on a log-log coordinate system, and linear regression is performed on the data points. In the log-log coordinate system, the power function relationship between the PPV and the proportional distance is converted into a linear relationship. The intercept of the fitted line is the logarithm of the site coefficient K, and the slope of the fitted line is the negative of the attenuation index α. The corrected site coefficient K is obtained by taking the exponent from the intercept of the fitted line, and the corrected attenuation index α is obtained by taking the negative of the slope of the fitted line. Substituting the corrected site coefficient K and the corrected attenuation index α, along with the allowable peak vibration velocity PPV and the distance R from the blast source to the monitoring point in the current blasting, into the algebraically transformed formula, the updated single-segment maximum charge threshold is calculated.

[0071] As the number of blasting excavations increases, the amount of accumulated field monitoring data gradually increases, the number of data points involved in the fitting increases, and the corrected site coefficient K and attenuation index α gradually approach the actual attenuation characteristic parameters under the actual geological conditions and blasting conditions of the tunnel project. The maximum charge threshold for a single section calculated based on the corrected parameters also gradually approaches the reasonable control value under actual engineering conditions, realizing the dynamic updating of blasting strength control parameters with the construction process.

[0072] In one or more embodiments of the present invention, applying the support structure includes: spraying a first concrete layer on the surface of the excavation face; erecting a steel arch frame on the surface of the first concrete layer along the tunnel excavation outline; drilling anchor bolt holes and installing system anchor bolts at the arch crown, arch shoulder, arch waist, and arch foot positions of the excavation face; spraying a second concrete layer to the designed thickness after the steel arch frame and system anchor bolts are installed; after the second concrete layer is sprayed, drilling locking anchor bolt holes and installing locking anchor bolts at the arch foot positions; the installation direction of the locking anchor bolts is inclined at an angle to the horizontal plane; and the number of locking anchor bolts installed at the same arch foot position is determined according to the degree of fracture of the surrounding rock at the arch foot.

[0073] After each blasting excavation, a support structure is applied to the excavation face. Following the completion of blasting excavation and the removal of ventilation, dust, and unstable rock, a concrete spraying machine is used to spray the mixed concrete mixture onto the excavation face surface using high-pressure airflow. During spraying, the spray gun maintains a consistent spraying distance from the excavation face surface and moves uniformly upwards circumferentially along the excavation face. The concrete mixture is evenly accumulated on the excavation face surface, gradually thickening to form the first concrete layer covering the entire excavation face. This first concrete layer seals the excavation face surface, filling in any uneven areas and preventing loose surface rock from detaching during subsequent construction. It also provides a bearing surface in close contact with the surrounding rock for the erection of the steel arch frame.

[0074] After the first concrete layer has set, a steel arch frame is erected on the surface of the first concrete layer along the tunnel excavation outline. The steel arch frame is assembled from multiple arc-shaped segments pre-fabricated outside the tunnel. During erection, the arch crown segment is erected first, followed by the arch shoulder segment and arch waist segment in sequence. The segments are fastened together with connecting plates and high-strength bolts to form a closed ring-shaped steel support structure along the excavation outline. The gap between the outer edge of the steel arch frame and the surface of the first concrete layer is tightly filled with wedge-shaped pads to ensure reliable force transmission contact between the steel arch frame and the surface of the first concrete layer.

[0075] After the steel arch frame is erected, systematic anchor bolt holes are drilled into the surrounding rock at the arch crown, arch shoulder, arch waist, and arch foot positions on the excavation face. Drilling is carried out using a rock drill, and rock debris is removed from the holes using high-pressure air after drilling. The anchor bolts are inserted into the anchor bolt holes, and anchoring agent is injected into the circumferential gap between the anchor bolt and the hole wall. The anchoring agent is injected from the bottom of the hole towards the opening until it overflows from the opening. After the anchoring agent reaches its set strength, washers and nuts are installed sequentially on the exposed ends of the anchor bolts, and pre-tightening force is applied. The systematic anchor bolts are drilled radially into the surrounding rock along the excavation face, connecting the shallow loosened zone of the surrounding rock with the deep stable rock mass, thereby improving the overall bearing capacity of the surrounding rock.

[0076] After the system anchor bolts are installed, a second layer of shotcrete is applied. This second layer covers the outer edge of the steel arch and the system anchor bolt pads, spraying to the designed thickness. During spraying, the speed of the spray gun and its distance from the sprayed surface are controlled. The concrete mix evenly wraps around the steel arch and fills the gaps between the steel arch and the surrounding rock, forming a composite support structure where the steel arch and the concrete layer work together. Once the second concrete layer reaches the designed thickness, the steel arch and system anchor bolts are encased within the concrete layer. The overall load-bearing capacity of the support structure is jointly borne by the compressive stress of the concrete layer, the bending stress of the steel arch, and the tensile stress of the system anchor bolts.

[0077] After the second concrete layer is sprayed, anchor bolt holes are drilled into the surrounding rock at the arch foot location, and anchor bolts are installed. The drilling direction of the anchor bolts is at an angle to the horizontal plane, pointing inwards into the surrounding rock. The borehole passes through the first concrete layer from the surface of the second concrete layer at the arch foot location into the surrounding rock. The installation sequence of the anchor bolts is the same as that of the system anchor bolts. A washer and nut are installed on the exposed end, and a preload is applied. The anchor bolts provide radial anchoring force to the arch foot, limiting the displacement and deformation of the arch foot under subsequent blasting disturbances and inhibiting the further expansion of internal fissures in the surrounding rock at the arch foot location. The number of anchor bolt holes drilled at the same arch foot location is determined according to the degree of fissure development in the surrounding rock at that location. During drilling, the degree of rock cuttings fragmentation and drilling speed are observed. When the degree of rock cuttings fragmentation is high or the drilling speed deviates from the normal range, the number of holes is increased. When the rock cuttings are intact fragments and the drilling speed is within the normal range, the basic number of holes is maintained. The anchor bolt holes are evenly distributed circumferentially on the surface of the second concrete layer at the arch foot location.

[0078] In one or more embodiments of the present invention, before grouting and filling the original fissures in the surrounding rock mass of the tunnel, the method further includes: sampling the surrounding rock of the tunnel and conducting SHPB dynamic impact tests to obtain the dynamic stress-strain curves and failure modes of the surrounding rock under different impact velocities, establishing a numerical model with the same size as the SHPB test device, and adjusting the parameters of the RHT dynamic constitutive model with the dynamic stress-strain curves and failure modes as calibration targets.

[0079] Furthermore, after adjusting the parameters of the RHT dynamic constitutive model, the process also includes: establishing a numerical model for cyclic blasting excavation of a small-clearance tunnel based on the preliminary borehole design scheme; substituting the calibrated RHT dynamic constitutive model parameters into the numerical model; simulating multiple blasting excavation processes of the lead tunnel and the follow tunnel; extracting damage variable cloud maps and damage depth data; identifying locations where damage variables exceed preset damage thresholds and damage depths exceed preset depth thresholds as damage-sensitive areas requiring pre-reinforcement; and adjusting the charge structure and detonation sequence of boreholes at the center of the tunnel face and at the excavation contour based on the simulation results.

[0080] Before grouting and filling the original fissures in the surrounding rock mass of the tunnel, surrounding rock samples were taken from the tunnel construction site and processed into cylindrical specimens with diameters and heights conforming to the standard dimensions of the split Hopkinson bar test. A dynamic impact test was conducted on the surrounding rock specimens using a split Hopkinson bar test system. This system consists of an impact bar, an incident bar, a transmission bar, and an absorbing bar. The material density, elastic modulus, and longitudinal wave velocity of the bars were consistent with the actual parameters of the test apparatus. The surrounding rock specimen was clamped between the incident bar and the transmission bar. A pulse shaper was attached to the impact end face of the incident bar to smooth the incident wave, and vacuum grease was applied to both end faces of the surrounding rock specimen to reduce end-face friction.

[0081] By adjusting the initial velocity of the impact rod, impact loading at different rates was applied to the surrounding rock specimen. After the impact rod struck the incident rod, a compression wave was generated in the incident rod, which propagated along the incident rod towards the specimen. When the compression wave reached the contact surface between the incident rod and the specimen, due to the difference in wave impedance between the specimen material and the rod material, part of the compression wave was reflected back to the incident rod to form a reflected wave, while the other part passed through the specimen and entered the transmission rod to form a transmitted wave. Strain gauges attached to the middle of the incident rod and the transmission rod recorded the incident wave signal, the reflected wave signal, and the transmitted wave signal, respectively. Based on the one-dimensional stress wave propagation theory, the dynamic strain of the specimen was calculated from the reflected wave signal, the dynamic stress of the specimen was calculated from the transmitted wave signal, and the strain rate of the specimen was calculated from the incident wave signal. The dynamic stress and dynamic strain calculated at the same impact velocity were plotted to obtain the dynamic stress-strain curves of the surrounding rock at different impact velocities. After each impact test, the broken fragments of the surrounding rock specimen were collected, and the size distribution characteristics and fracture surface morphology of the fragments were recorded as the failure mode of the surrounding rock specimen at that impact velocity.

[0082] To obtain constitutive model parameters that reflect the dynamic mechanical behavior of surrounding rock under high strain rate loading conditions during blasting, a three-dimensional numerical model with dimensions identical to the split Hopkinson bar test setup was established. This numerical model includes an impact bar, an incident bar, a transmission bar, an absorbing bar, and a surrounding rock specimen, with the geometric dimensions of the surrounding rock specimen being identical to those of the experimental specimen. The material density, elastic modulus, longitudinal wave velocity, and geometric dimensions of each bar are consistent with the actual parameters of the test setup. A simulated shaper of the same size as the pulse shaper used in the experiment is placed at the center of the incident bar's end face. A stress wave is generated by applying the same initial velocity as in the experiment to the impact bar. After propagating along the incident bar to the specimen, part of this stress wave enters the transmission bar, and the other part is reflected back to the incident bar, mirroring the stress wave propagation process in the experiment.

[0083] In the numerical model, the surrounding rock specimens are described using the RHT dynamic constitutive model. The basic mechanical parameters required for the RHT dynamic constitutive model are obtained from laboratory rock mechanics tests, including density, Poisson's ratio, uniaxial compressive strength, uniaxial tensile strength, and elastic modulus. The RHT dynamic constitutive model introduces three limit surfaces: the elastic limit surface, the failure surface, and the residual strength surface. The elastic limit surface describes the stress threshold value at which the surrounding rock transitions from the elastic stage to the plastic stage. The failure surface describes the peak strength that the surrounding rock can reach under different hydrostatic pressure conditions. The residual strength surface describes the residual bearing capacity retained by the surrounding rock after failure. The damage variable D in the RHT dynamic constitutive model is defined by the ratio of accumulated plastic strain to failure strain. When the stress state reaches the limit strength of the failure surface, damage caused by plastic strain begins to accumulate. The value of the damage variable D ranges from zero at the point of no damage to its maximum value at the point of complete failure. The RHT dynamic constitutive model describes the rate dependence of rock strength as strain rate increases through compressive strain rate exponent and tensile strain rate exponent, and describes the nonlinear relationship between pressure and volumetric strain during the gradual crushing and compaction of pores under pressure through the p-α equation of state.

[0084] Using the dynamic stress-strain curves and specimen failure modes obtained from the split Hopkinson bar dynamic impact test as calibration targets, stress-strain output curves were extracted from the numerical model at the same locations as the experimental strain gauges. The stress-strain curves output by the numerical model were then compared with those obtained from the experiment. The comparison included the slope of the elastic loading segment of the curve, the peak stress value, the strain value corresponding to the peak stress, the post-peak stress drop rate, and the residual stress level. Simultaneously, the distribution of failure elements in the surrounding rock specimen after impact loading in the numerical model was compared with the specimen fragment size distribution characteristics and fracture surface morphology recorded in the experiment. The failure surface parameters, tensile-compressive meridian ratio, damage parameters, and residual strength surface parameters in the RHT dynamic constitutive model were adjusted using a trial-and-error method. After these parameter adjustments, the numerical model was rerun, and new stress-strain output curves and failure modes were extracted and compared with the experimental results again. Repeat the above adjustment and comparison process. When the deviation between the stress-strain curve output by the numerical model and the experimental curve in the above comparison items is within the preset deviation range, and the specimen fragmentation degree and fragment distribution characteristics simulated by the numerical model are consistent with the failure mode recorded in the test, the RHT dynamic constitutive model parameters at this time are the calibrated surrounding rock dynamic constitutive parameters, which can be used to establish the numerical model for subsequent small-clearance tunnel cyclic blasting excavation.

[0085] After calibrating the RHT dynamic constitutive model parameters, a numerical model for cyclic blasting excavation of a tunnel with small clearances was established based on the preliminary borehole design scheme. The preliminary borehole design scheme included the charge structure type and detonation sequence of boreholes at the center of the tunnel face and at the excavation outline. The geometric dimensions of the numerical model were determined based on the actual tunnel cross-sectional dimensions, the clearance between the preceding and following tunnels, and the single-cycle advance. The surrounding rock of the tunnel was divided into multiple unit regions in the numerical model, and the calibrated RHT dynamic constitutive model parameters were assigned as the mechanical parameters of the surrounding rock units. The boundaries of the numerical model were set as non-reflective boundaries to allow blasting stress waves to be transmitted at the boundaries without reflection and superposition, simulating the stress wave propagation conditions of actual surrounding rock in a semi-infinite space.

[0086] Following the design and construction sequence of tunnels with small clearances, numerical simulations of multiple blasting excavations were conducted for both the initial and subsequent tunnels. The simulation for the initial tunnel began at the starting position of the tunnel face and advanced sequentially in a single-cycle advance. The simulation for the subsequent tunnel began after the distance between the tunnel faces met the set offset condition. After each blasting excavation simulation, the damage variable values ​​of each unit in the surrounding rock were saved in the numerical model. The damage variable values ​​of each unit on the same cross-section were extracted radially outward along the tunnel excavation profile, forming a damage variable cloud map of that cross-section after the completion of that blasting excavation. From the damage variable cloud map, the distribution range of surrounding rock units whose damage variables exceeded the preset damage threshold was statistically analyzed along the normal direction of the tunnel excavation profile. The maximum extension length of this distribution range along the normal direction was determined as the damage depth. When the damage variable at a certain location on a certain cross-section exceeded the preset damage threshold and the damage depth exceeded the preset depth threshold, that location was identified as a damage-sensitive area requiring pre-reinforcement.

[0087] After identifying the damage-sensitive areas based on the simulation results, the preliminary borehole design scheme is adjusted accordingly. If the simulation results show that the damage depth at the arch shoulder or arch foot exceeds the preset depth threshold, the charging structure of the boreholes at the corresponding excavation outline position is adjusted. The continuous charging structure is changed to an intermittent charging structure, or the number of intervals in the intermittent charging structure is increased to reduce the proportion of energy transferred to the remaining surrounding rock. If the simulation results show that the free surface formation effect in the central area of ​​the working face is insufficient and unbroken rock mass remains after the cut hole blasting, the charging structure parameters of the boreholes at the central position of the working face are adjusted. The proportion of the charge in the concentrated charge is increased to the total charge, or the difference between the delay time of the cut hole and the delay time of the subsequent auxiliary holes in the detonation sequence is adjusted so that the free surface formed after the cut hole blasting has sufficient time to fully develop. Under the guidance of this free surface, the blasting energy of the subsequent boreholes is concentrated and released towards the rock mass to be excavated, improving the directional utilization of blasting energy.

[0088] In one or more embodiments of the present invention, the method further includes: obtaining the variation curves of surrounding rock deformation and support structure stress with excavation cycle through cyclic blasting numerical simulation, determining the construction parameters of the support structure based on the variation curves, and determining the specifications of the support structure so that the bearing capacity of the support structure is greater than the simulated surrounding rock pressure.

[0089] After establishing a numerical model for the cyclic blasting excavation of small-clearance tunnels and determining the damage-sensitive areas, this numerical model was used to simulate the entire process of multiple cyclic blasting excavations of the lead tunnel and the follow tunnel. After each blasting excavation simulation, the radial deformation of the surrounding rock at various characteristic locations around the excavation face and the stress values ​​of each component of the constructed support structure were extracted. The locations for extracting the radial deformation of the surrounding rock included the arch crown, arch shoulder, arch waist, and arch foot of the excavation face. The stress values ​​of the support structure components included the axial force and bending moment of each segment of the steel arch frame, the axial force of the system anchor bolts, and the axial force of the locking anchor bolts. The radial deformation at the same characteristic location after each blasting excavation was connected into a curve according to the excavation cycle sequence to obtain the curve of the change in surrounding rock deformation with the excavation cycle. Similarly, the stress values ​​of the same support component after each blasting excavation were connected into a curve according to the excavation cycle sequence to obtain the curve of the change in the stress of the support structure with the excavation cycle.

[0090] The curve showing the change in surrounding rock deformation with excavation cycles reveals a trend of gradual development in surrounding rock deformation as the tunnel face advances and subsequent blasting disturbances accumulate. Similarly, the curve showing the change in support structure stress with excavation cycles reveals a trend of gradual increase in support structure load as surrounding rock deformation increases. Based on the rate of deformation growth and the final cumulative deformation in the surrounding rock deformation curve, the required support resistance at each characteristic location is determined. Furthermore, based on the maximum stress value of each component and the distribution range of the stress growth segment in the support structure stress curve, the required bearing capacity of each component of the support structure is determined.

[0091] Based on the aforementioned curves, the construction parameters for the support structure are determined. These parameters include the cross-sectional specifications of the steel arch frame, the length and spacing of the system anchors, the design thickness of the shotcrete layer, and the length and number of locking anchors. The cross-sectional specifications of the steel arch frame are determined based on the maximum radial deformation of the surrounding rock at the arch crown, ensuring that the cross-sectional bearing capacity of the steel arch frame under the corresponding axial force and bending moment combination is greater than the simulated maximum stress value. The length and spacing of the system anchors are determined based on the maximum radial deformation of the surrounding rock at the arch shoulder and arch waist. The length of the system anchors ensures that their anchoring sections extend into the stable rock mass outside the loosened zone, and the spacing of the system anchors ensures that the effective bearing range of adjacent anchors overlaps in the circumferential direction. The design thickness of the first and second concrete layers is determined based on the slope of the deformation growth segment in the surrounding rock deformation curve. The total thickness of the shotcrete layer ensures that the compressive bearing capacity of the concrete layer under the corresponding surrounding rock pressure is greater than the simulated surrounding rock pressure. The length and number of anchor bolts are determined based on the maximum radial deformation of the surrounding rock at the arch foot and the maximum axial force in the anchor bolt stress variation curve. The bearing capacity of the anchor bolts is greater than the maximum axial force value of the anchor bolts obtained from the simulation. By using the above method, the overall bearing capacity of the support structure is made greater than the surrounding rock pressure obtained from the numerical simulation, ensuring that the support structure has sufficient bearing margin throughout the entire cyclic blasting construction process.

[0092] This invention achieves proactive control of surrounding rock damage in cyclic blasting construction of tunnels with small clearances. Grouting pretreatment fills the original fissures in the rock mass and blocks groundwater seepage channels, fundamentally eliminating the promoting effect of hydrodynamic pressure caused by blasting stress waves on fissure propagation. Based on the difference in energy transmission paths between blast holes at the center of the tunnel face and those at the excavation outline, continuous and intermittent charging structures are respectively set up and coordinated with a center-first detonation sequence, ensuring blasting energy is directionally transmitted towards the rock mass to be excavated, reducing energy dissipation towards the remaining surrounding rock. Pre-reinforcement is applied at the arch shoulder, arch foot, and interbedded rock areas to enhance the initial damage resistance of damage-sensitive areas. For each blasting excavation, the maximum charge threshold for a single section is determined based on the clearance between the preceding and following tunnels, the surrounding rock grade, and the allowable peak vibration velocity, and the actual charge is controlled to not exceed this threshold. Simultaneously, support structures are promptly applied after each blasting excavation to inhibit further propagation of surrounding rock fissures. The above four steps run through the entire construction process, and progressive intervention is carried out from four dimensions: water-induced conditions, energy sources, structural weaknesses and propagation paths. This forms a complete active control closed loop for the damage to the surrounding rock caused by cyclic blasting in tunnels with small clearance, which significantly reduces the cumulative damage to the surrounding rock caused by cyclic blasting.

[0093] The above is the overall concept of the present invention. For ease of understanding, the present invention also provides the following embodiments:

[0094] In the construction of tunnels with small clearances, engineering geological and blasting design parameters are obtained, including the tunnel surrounding rock type, surrounding rock grade, tunnel cross-sectional dimensions, left and right tunnel clearance, single-cycle advance, construction sequence, borehole layout, borehole diameter, charge diameter, borehole spacing, charge structure, delay section number, and explosive type. The tunnel surrounding rock is moderately weathered granite, with a surrounding rock grade of Class III. The drill-and-blast method is used for construction, with a borehole diameter of 42mm and a single-cycle advance of 3.0m. The boreholes are arranged as cut holes, auxiliary holes, collapse holes, buffer holes, peripheral holes, and bottom holes, and multiple delay sections are set.

[0095] Before tunnel construction, grouting is carried out on the surrounding rock within the tunnel excavation area. The grouting material fills the original fissures in the rock mass, sealing the originally connected fissure network into isolated pores, thereby blocking the seepage channels of groundwater along the fissure surfaces inside the rock mass.

[0096] The layout of grouting holes extends outward along the tunnel excavation outline, and the final depth of the grouting holes extends beyond the excavation outline. The selection of grouting materials is determined comprehensively based on the degree of development of surrounding rock fissures and groundwater occurrence conditions. Grouting is terminated when the grouting pressure reaches the designed final pressure and the grouting volume per unit time remains below the preset termination threshold.

[0097] The mechanism by which grouting blocks groundwater seepage channels lies in eliminating the coupling effect between blasting stress waves and fissure water. In ungrouted surrounding rock, groundwater exists within rock fissures. When blasting stress waves propagate to the water-bearing fissure surface, the instantaneous high pressure caused by the stress waves acts on the fissure water, forcing it into the fissure tip and creating a water wedge effect. Simultaneously, the stress waves directly damage the rock mass framework. The water wedge effect and the stress wave damage effect mutually promote each other. Under the continuous action of the stress waves, the fissure water is repeatedly forced into the fissure tip, and the stress intensity factor at the fissure tip gradually increases. When it exceeds the fracture toughness of the rock mass, the fissure expands. The expanded fissure then provides a channel for more fissure water to enter, forming a positive feedback loop that accelerates damage.

[0098] After grouting, the grout fills the rock fissures and solidifies. The space occupied by the solidified grout in the fissures no longer has the conditions for the existence of free water. When the blasting stress wave propagates to the filled fissures, the carrier of the dynamic water pressure effect is eliminated because there is no free water medium in the fissures. The preconditions for the water wedge splitting effect no longer hold, and the positive feedback mechanism of water-induced damage acceleration is weakened from the root.

[0099] Before grouting and filling the original fissures in the surrounding rock mass of the tunnel, samples of the surrounding rock were taken from the tunnel construction site. These samples were then processed into cylindrical specimens with diameters and heights conforming to the standard dimensions of the split Hopkinson bar test. Dynamic impact tests were conducted on the surrounding rock specimens using a split Hopkinson bar test system. This system consists of an impact bar, an incident bar, a transmission bar, and an absorbing bar. The material density, elastic modulus, and longitudinal wave velocity of the bars were consistent with the actual parameters of the testing apparatus.

[0100] The surrounding rock specimen is sandwiched between the incident rod and the transmission rod. A pulse shaper is attached to the impact end face of the incident rod to smooth the incident wave, and vacuum grease is applied to both ends of the surrounding rock specimen to reduce end-face friction. Different impact loading rates are applied to the surrounding rock specimen by adjusting the initial velocity of the impact rod. After the impact rod strikes the incident rod, a compression wave is generated in the incident rod, which propagates along the incident rod towards the specimen. When the compression wave reaches the contact surface between the incident rod and the specimen, due to the difference in wave impedance between the specimen material and the rod material, part of the compression wave is reflected back to the incident rod to form a reflected wave, while the other part passes through the specimen and enters the transmission rod to form a transmitted wave. Strain gauges attached to the middle of the incident rod and the transmission rod record the incident wave signal, the reflected wave signal, and the transmitted wave signal, respectively. Based on the one-dimensional stress wave propagation theory, the dynamic strain of the specimen is calculated from the reflected wave signal, the dynamic stress is calculated from the transmitted wave signal, and the strain rate is calculated from the incident wave signal. The dynamic stress and dynamic strain calculated at the same impact velocity are plotted to obtain the dynamic stress-strain curves of the surrounding rock at different impact velocities. After each impact test, the broken fragments of the surrounding rock specimen are collected, and the size distribution characteristics and fracture surface morphology of the fragments are recorded as the failure mode of the surrounding rock specimen at that impact velocity.

[0101] To obtain constitutive model parameters that reflect the dynamic mechanical behavior of the surrounding rock under high strain rate loading conditions during blasting, a three-dimensional numerical model with dimensions identical to the split Hopkinson bar test setup was established. This numerical model includes an impact bar, an incident bar, a transmission bar, an absorbing bar, and a surrounding rock specimen, with the geometric dimensions of the surrounding rock specimen being exactly the same as those of the experimental specimen. The material density, elastic modulus, longitudinal wave velocity, and geometric dimensions of each bar are consistent with the actual parameters of the test setup. A simulated shaper of the same size as the pulse shaper used in the experiment is set at the center of the end face of the incident bar. By applying the same initial velocity as in the experiment to the impact bar, a stress wave is generated. After propagating along the incident bar to the specimen, part of the stress wave enters the transmission bar, and the other part is reflected back to the incident bar, which is the same propagation process of the stress wave in the experiment.

[0102] In the numerical model, the surrounding rock specimens are described using the RHT dynamic constitutive model. The basic mechanical parameters required for the RHT dynamic constitutive model are obtained from laboratory rock mechanics tests, including density, Poisson's ratio, uniaxial compressive strength, uniaxial tensile strength, and elastic modulus. The RHT dynamic constitutive model introduces three limit surfaces: the elastic limit surface, the failure surface, and the residual strength surface. The elastic limit surface describes the stress threshold value at which the surrounding rock transitions from the elastic stage to the plastic stage. The failure surface describes the peak strength that the surrounding rock can reach under different hydrostatic pressures. The residual strength surface describes the residual bearing capacity retained by the surrounding rock after failure. The damage variable D in the RHT dynamic constitutive model is defined by the ratio of cumulative plastic strain to failure strain. When the stress state reaches the limit strength of the failure surface, the damage caused by plastic strain begins to accumulate. The RHT dynamic constitutive model describes the rate dependence of rock strength as the strain rate increases through the compressive strain rate exponent and the tensile strain rate exponent, and describes the nonlinear relationship between pressure and volumetric strain during the gradual crushing and compaction of pores under pressure through the p-α equation of state.

[0103] Using the dynamic stress-strain curves and specimen failure modes obtained from the split Hopkinson bar dynamic impact test as calibration targets, stress-strain output curves were extracted from the numerical model at the same locations as the experimental strain gauges. The stress-strain curves output by the numerical model were compared with those obtained from the experiment. The comparison included the slope of the elastic loading segment of the curve, the peak stress value, the strain value corresponding to the peak stress, the stress drop rate after the peak, and the residual stress level. Simultaneously, the distribution of failure elements in the surrounding rock specimen after impact loading in the numerical model was compared with the specimen fragment size distribution characteristics and fracture surface morphology recorded in the experiment.

[0104] The failure surface parameters, tensile-compressive meridian ratio, damage parameters, and residual strength surface parameters in the RHT dynamic constitutive model were adjusted using a trial-and-error method. After these adjustments, the numerical model was run again to extract new stress-strain output curves and failure modes, which were then compared with the experimental results. This adjustment and comparison process was repeated. When the deviations between the stress-strain curves output by the numerical model and the experimental curves in the above comparison items were within the preset deviation range, and the specimen fragmentation degree and fragment distribution characteristics simulated by the numerical model tended to be consistent with the failure modes recorded in the experiments, the RHT dynamic constitutive model parameters at this point were considered the calibrated dynamic constitutive parameters of the surrounding rock.

[0105] After calibrating the RHT dynamic constitutive model parameters, a numerical model for cyclic blasting excavation of a tunnel with small clearances was established based on the preliminary borehole design scheme. The preliminary borehole design scheme included the charge structure type and detonation sequence of boreholes at the center of the tunnel face and at the excavation outline. The geometric dimensions of the numerical model were determined based on the actual tunnel cross-sectional dimensions, the clearance between the preceding and following tunnels, and the single-cycle advance. The surrounding rock of the tunnel was divided into multiple unit regions in the numerical model, and the calibrated RHT dynamic constitutive model parameters were assigned as the mechanical parameters of the surrounding rock units.

[0106] The numerical model's boundaries are set to non-reflective boundaries to allow blasting stress waves to transmit at the boundaries without reflection and superposition, simulating the actual stress wave propagation conditions in semi-infinite space. Following the design and construction sequence of small-clearance tunnels, numerical simulations of multiple blasting excavations were conducted for both the pilot tunnel and the subsequent tunnel. The simulation of the pilot tunnel's blasting excavation began at the starting position of the pilot tunnel face and progressed sequentially in a single cycle. The simulation of the subsequent tunnel's blasting excavation began after the distance between the subsequent tunnel face and the pilot tunnel face met the set offset condition.

[0107] After each blasting excavation simulation, the damage variable values ​​of each unit in the surrounding rock in the numerical model are saved. The damage variable values ​​of each unit on the same cross-section are extracted radially outward along the tunnel excavation profile to form a damage variable cloud map of that cross-section after the completion of the blasting excavation. From the damage variable cloud map, the distribution range of surrounding rock units whose damage variables exceed the preset damage threshold is statistically analyzed along the normal direction of the tunnel excavation profile. The maximum extension length of this distribution range along the normal direction is determined as the damage depth. When the damage variable at a certain location on a certain cross-section exceeds the preset damage threshold and the damage depth exceeds the preset depth threshold, that location is determined as a damage-sensitive area requiring pre-reinforcement.

[0108] After identifying the damage-sensitive areas based on the simulation results, the preliminary borehole design scheme is adjusted accordingly. If the simulation results show that the damage depth at the arch shoulder or arch foot exceeds the preset depth threshold, the charging structure of the boreholes at the corresponding excavation outline position is adjusted, changing the continuous charging structure to an intermittent charging structure and increasing the number of intervals in the intermittent charging structure to reduce the proportion of energy transferred to the remaining surrounding rock. If the simulation results show that the free surface formation effect in the central area of ​​the working face is insufficient and unbroken rock mass remains after the cut hole blasting, the charging structure parameters of the boreholes at the central position of the working face are adjusted, increasing the proportion of the charge in the concentrated charge to the total charge, or adjusting the difference between the delay time of the cut hole and the delay time of the subsequent auxiliary holes in the detonation sequence, so that the free surface formed after the cut hole blasting has sufficient time to fully develop, and the blasting energy of the subsequent boreholes is concentrated and released towards the rock mass to be excavated under the guidance of the free surface, thereby improving the directional utilization of blasting energy.

[0109] The layout of blast holes and the design of the charging structure are carried out at the tunnel face. Based on their location on the tunnel face, the blast holes are divided into two categories: blast holes at the center of the tunnel face and blast holes along the excavation outline.

[0110] The blast hole at the center of the working face is located in the middle region of the working face. This type of blast hole is surrounded by intact rock mass, with only one free face facing the working face. The rock mass exerts a significant clamping force on this type of blast hole. After the explosive inside the blast hole detonates, it needs to overcome the clamping force of the rock mass and break apart to create a cavity facing the working face, thus creating a new free face. The blast hole at the center of the working face uses a continuous charge structure. A continuous charge structure means that explosive charges are continuously loaded along the axial direction of the blast hole without any spacer between the charges. During blasting, the charges continuously detonate along the axial direction of the blast hole, and the energy is concentrated and released radially, making the pressure exerted by the blast on the surrounding rock mass more concentrated.

[0111] The blast holes at the excavation outline location are situated in the area surrounding the working face. One side of these blast holes is the fractured rock mass to be excavated, while the other side is the surrounding rock to be preserved. After the explosives detonate, the energy is simultaneously transmitted in both directions: the side to be excavated and the side to be preserved. The blast holes at the excavation outline location employ a spaced-charge structure. This structure involves segmented loading of explosive cartridges along the axial direction of the blast hole, with spacers between adjacent cartridges. These spacers are not filled with explosives; the spacer medium is air or a spacer made of inert material. During blasting, the explosive cartridges detonate segmentally along the axial direction of the blast hole. The radial distribution of blast energy along the blast hole is more uniform due to the presence of the spacers, resulting in a more stable blast pressure distribution along the excavation outline surface and a reduced proportion of energy transmitted towards the preserved surrounding rock.

[0112] After the charge structure is determined, the detonation sequence is set so that the boreholes at the center of the tunnel face are detonated before the boreholes at the excavation outline. Among the boreholes at the center of the tunnel face, the cut hole is the first to detonate. The cut hole is located in the central area of ​​the tunnel face. After the cut hole detonates, it creates a cavity facing the tunnel face in the rock mass in the central area of ​​the tunnel face. The surface of this cavity is the new free surface that can be used for subsequent borehole blasting.

[0113] When the blasting stress wave propagates to the free surface, it is reflected. The compressive wave reflected from the free surface is unloaded and transformed into a tensile wave that propagates into the rock mass. The tensile wave superimposes with the incident compressive wave, forming a tensile stress zone near the free surface. When the tensile stress exceeds the tensile strength of the rock, the rock undergoes tensile failure in this zone. After the cut hole is detonated, the auxiliary hole at the center of the tunnel face is detonated. The auxiliary hole is located between the cut hole and the blast hole at the excavation outline. It detonates under the condition that the free surface created by the cut hole has already been formed. The blasting energy of the auxiliary hole is released directionally towards the free surface created by the cut hole, further expanding the range of the free surface. After the auxiliary hole is detonated, the peripheral holes at the excavation outline are detonated. The peripheral holes are adjacent to the tunnel excavation design outline. They detonate under the condition of the expanded free surface. The blasting energy is released uniformly along the excavation outline, completing the formation of the tunnel cross-section outline.

[0114] After the peripheral boreholes are detonated, the bottom boreholes are detonated. Located at the bottom of the tunnel face, these boreholes are detonated after the cross-sectional profile has been largely formed, thus breaking up the rock mass at the bottom of the tunnel face and forming the tunnel floor profile. The delay time between adjacent sections is set according to the staggered propagation conditions of the blasting vibration waves. The delay time is determined by obtaining the distance from each section's borehole to the surrounding rock monitoring point, calculating the arrival time of the blasting seismic waves from the blast source to the monitoring point based on the propagation velocity of seismic waves in the surrounding rock, and using the difference in arrival times between adjacent sections as the minimum delay time. This ensures that the peak value of the seismic wave from the first detonated section passes the monitoring point before the peak value of the seismic wave from the subsequent detonated section arrives at the same monitoring point, avoiding the simultaneous arrival of multiple seismic wave peaks at the same location in the surrounding rock, which would cause vibration superposition and amplification.

[0115] When calculating the equivalent blasting load, the peak pressure on the borehole wall is calculated based on the explosive density, detonation velocity, borehole diameter, charge diameter, radial decoupling coefficient, axial decoupling coefficient, and number of boreholes. For boreholes detonated simultaneously within the same delay period, the blasting pressure acting on each borehole wall is converted into blasting pressure applied to the equivalent boundary. The equivalent blasting load is calculated separately for cut holes, auxiliary holes, caving holes, buffer holes, peripheral holes, and bottom holes. A triangular load function is used to approximate the blasting pressure time history. The triangular load function includes a rising phase and a decay phase. The rising time is calculated from the moment the blasting pressure begins to act on the borehole wall until the pressure reaches its peak value, and the duration is calculated from the moment the pressure rises until the pressure decays to a negligible value. The values ​​of the rising time and duration are determined based on the explosive type and charge structure parameters. The equivalent blasting load is applied to the corresponding excavation zone boundary according to the borehole type.

[0116] After completing the design of the borehole charge structure and setting the detonation sequence, pre-reinforcement was implemented on the arch shoulder, arch foot, and interbedded rock areas of the surrounding rock. The arch shoulder is the transition arc section between the arch and sidewall of the tunnel excavation outline, and the arch foot is the junction of the sidewall and bottom slab of the tunnel excavation outline. Both of these are geometric inflection points of the excavation outline. When the blasting stress wave propagates to the geometric inflection point, the abrupt curvature change of the wavefront at the inflection point causes diffraction and reflection of the stress wave, making the dynamic stress concentration in the surrounding rock near the inflection point higher than that in the straight outline area. Under the repeated loading of multiple blasting cycles, the damage accumulation rate and degree at the geometric inflection point are greater than those in other outline areas. The interbedded rock area is the rock column between the first and second tunnels. During construction, it is subjected to the alternating and superimposed effects of blasting disturbances from both directions. The bidirectional disturbance causes the micro-cracks inside the interbedded rock to expand alternately in both directions, making the damage accumulation path more complex and the damage evolution rate faster. Pre-reinforcement improves the initial damage resistance of the three types of damage-sensitive areas by applying reinforcement structures in advance.

[0117] After pre-reinforcement of the arch shoulder, arch foot, and interbedded rock areas, blasting excavation is carried out. Before each blasting excavation, the maximum charge threshold for a single segment is determined. Depending on whether the blast is the first blast or a subsequent cycle, different parameter sources are used for threshold calculation. In the empirical relationship between peak blast velocity and charge amount, the peak blast velocity increases with the increase of the single-segment charge amount and decreases with the increase of the distance from the blast source to the monitoring point. To solve for the maximum charge threshold for a single segment, this relationship is algebraically transformed, expressing the single-segment charge amount as a dependent variable jointly determined by the peak blast velocity, site coefficient, attenuation index, and distance from the blast source to the monitoring point. Substituting the allowable peak blast velocity as the value of the peak blast velocity into the algebraically transformed relationship, the calculated single-segment charge amount is the maximum charge threshold for a single segment.

[0118] For the initial blasting excavation, the site coefficient and attenuation index are selected from the empirical values ​​of blasting vibration attenuation parameters for the same level of surrounding rock as specified in the blasting safety regulations. Using the tunnel surrounding rock level as an index, the suggested values ​​for the site coefficient and attenuation index corresponding to that rock level are retrieved from the empirical value table. For subsequent blasting excavations after the initial blast, the site coefficient and attenuation index are fitted and corrected using the on-site blasting vibration monitoring data from each completed blast. The actual charge amount and distance from the blast source to the monitoring point in each segment of each blast are converted into proportional distances. The proportional distances and the corresponding measured peak vibration velocities are plotted on a double logarithmic coordinate system. Linear regression fitting is performed on the data points, and the corrected site coefficient is obtained by taking the exponent from the intercept of the fitted line. The corrected attenuation index is obtained by taking the negative of the slope of the fitted line. The actual charge amount of each delay segment in this blasting design is compared with the maximum charge amount threshold for a single segment. If the charge amount of a certain delay segment exceeds the threshold, the charge amount of the delay segment is reduced to below the threshold, or the delay segment is split into two or more delay segments so that the charge amount of each segment after splitting does not exceed the maximum charge amount threshold for a single segment.

[0119] After each blasting excavation, a support structure is applied to the excavation face. After the blasting excavation is completed and ventilation, dust removal, and removal of unstable rocks are finished, a concrete spraying machine is used to spray the mixed concrete mixture onto the surface of the excavation face using a high-pressure airflow. During spraying, the spray gun maintains a spraying distance from the surface of the excavation face and moves at a uniform speed from bottom to top along the circumference of the excavation face. The concrete mixture is evenly accumulated on the surface of the excavation face and gradually thickens, forming the first concrete layer covering the entire excavation face.

[0120] The first concrete layer seals the surface of the excavation face, fills the local uneven areas of the excavation face, prevents loose surface rock blocks from falling off during subsequent construction, and provides a pressure-bearing base surface in close contact with the surrounding rock for the erection of the steel arch frame.

[0121] After the first concrete layer has set, a steel arch frame is erected on the surface of the first concrete layer along the tunnel excavation outline. The steel arch frame is composed of multiple arc-shaped segments pre-processed outside the tunnel. During erection, the arch top segment is erected first, followed by the arch shoulder segment and arch waist segment in sequence. The segments are fastened together by connecting plates and high-strength bolts to form a closed ring-shaped steel support structure along the excavation outline. The gap between the outer edge of the steel arch frame and the surface of the first concrete layer is tightly filled with wedge-shaped pads. After the steel arch frame is erected, system anchor bolt holes are drilled into the surrounding rock at the arch crown, arch shoulder, arch waist, and arch foot positions on the excavation face. Drilling is carried out using a rock drill. After drilling, rock debris is removed from the hole with high-pressure air. After the anchor bolt body is inserted into the anchor bolt hole, anchoring agent is injected into the circumferential gap between the anchor bolt body and the hole wall. The anchoring agent is injected from the bottom of the hole towards the opening until it overflows from the opening. After the anchoring agent reaches the setting strength, washers and nuts are installed in sequence at the exposed end of the anchor bolt body and pre-tightening force is applied to connect the shallow loose ring of the surrounding rock with the deep stable rock mass.

[0122] After the system anchor bolts are installed, a second layer of concrete is sprayed. The second layer of concrete covers the outer edge of the steel arch frame and the system anchor bolt pads, and is sprayed to the designed thickness. During the spraying process, the moving speed of the spray gun and the distance between it and the sprayed surface are controlled. The concrete mix evenly wraps the steel arch frame and fills the gap between the steel arch frame and the surrounding rock, forming a composite support structure in which the steel arch frame and the concrete layer work together. The overall bearing capacity of the support structure is jointly borne by the compression of the concrete layer, the bending of the steel arch frame, and the tension of the system anchor bolts.

[0123] After the second concrete layer is sprayed, anchor bolt holes are drilled into the surrounding rock at the arch foot location, and anchor bolts are installed. The drilling direction of the anchor bolts is at an angle to the horizontal plane, pointing inwards into the surrounding rock. The boreholes pass through the first concrete layer from the surface of the second concrete layer at the arch foot location into the surrounding rock. The installation sequence of the anchor bolts is the same as that of the system anchor bolts. Washers and nuts are installed on the exposed ends, and pre-tightening force is applied to provide radial anchoring force to the arch foot, limiting the displacement and deformation of the arch foot under subsequent blasting disturbance. The number of anchor bolt holes drilled at the same arch foot location is determined according to the degree of fissure development in the surrounding rock at that location. During drilling, the degree of rock cuttings fragmentation and drilling speed are observed. If the degree of rock cuttings fragmentation is high or the drilling speed deviates from the normal range, the number of holes is increased. If the rock cuttings are intact fragments and the drilling speed is within the normal range, the basic number of holes is maintained. The anchor bolt holes are evenly distributed circumferentially on the surface of the second concrete layer at the arch foot location.

[0124] During cyclic blasting excavation, numerical simulations of cyclic blasting are used to obtain curves showing the changes in surrounding rock deformation and support structure stress with each excavation cycle. After each blasting excavation simulation, the radial deformation at various characteristic locations of the surrounding rock around the excavation face and the stress values ​​of each component of the constructed support structure are extracted. The locations for extracting the radial deformation of the surrounding rock include the arch crown, arch shoulder, arch waist, and arch foot of the excavation face. The stress values ​​of the support structure components include the axial force and bending moment of each segment of the steel arch frame, the axial force of the system anchor bolts, and the axial force of the locking anchor bolts. By connecting the radial deformation at the same characteristic location after each blasting excavation cycle in the order of the excavation cycles, a curve is obtained showing the change in surrounding rock deformation with each excavation cycle. Similarly, by connecting the stress values ​​of the same support component after each blasting excavation cycle in the order of the excavation cycles, a curve is obtained showing the change in support structure stress with each excavation cycle.

[0125] The required support resistance at each characteristic location is determined based on the rate of deformation growth and the final cumulative deformation in the surrounding rock deformation curve. The required bearing capacity of each component of the support structure is determined based on the maximum stress value and the distribution range of the stress growth segment in the stress variation curve of the support structure. The cross-sectional specifications of the steel arch frame are determined based on the maximum radial deformation of the surrounding rock at the arch crown, ensuring that the cross-sectional bearing capacity of the steel arch frame under the corresponding axial force and bending moment combination is greater than the simulated maximum stress value. The length and spacing of the system anchor bolts are determined based on the maximum radial deformation of the surrounding rock at the arch shoulder and arch waist. The length of the system anchor bolts ensures that their anchoring sections extend into the stable rock mass outside the loosened zone of the surrounding rock. The spacing of the system anchor bolts ensures that the effective bearing range of adjacent anchor bolts overlaps in the circumferential direction. The total thickness of the shotcrete layer is determined based on the slope of the deformation growth segment in the surrounding rock deformation curve, ensuring that the compressive bearing capacity of the concrete layer under the corresponding surrounding rock pressure is greater than the simulated surrounding rock pressure. The length and number of the locking anchor bolts are determined based on the maximum radial deformation of the surrounding rock at the arch foot position and the maximum axial force in the locking anchor bolt force variation curve. The bearing capacity of the locking anchor bolts is greater than the simulated maximum axial force value of the locking anchor bolts.

[0126] Monitoring areas were established at typical cross-sections of the pilot and follow-up tunnels. The arch crown, left and right abutments, left and right arch waists, left and right arch feet, and the interbedded rock area were selected as characteristic locations. For the interbedded rock, multiple monitoring points with concentrated damage were further selected. The vibration velocity time histories of each monitoring point under different delay blasting stages were extracted to obtain the peak vibration velocity. The proportional distance was calculated based on the blast source distance and single-stage charge, establishing the attenuation relationship between the peak vibration velocity and the proportional distance. The type of blast hole that contributes significantly to the vibration response of the interbedded rock was determined based on the peak vibration velocity generated by blasting at different delay stages.

[0127] Based on the damage variable output results in the RHT model, damage variable cloud maps were extracted from typical cross-sections after each excavation cycle. Areas where the damage variable reached a set threshold were designated as damage zones. Damage depth was statistically analyzed along typical directions. The changes in damage depth at the crown, shoulder, waist, foot, and interbedded rock were compared after different excavation cycles to obtain the cumulative evolution law of surrounding rock damage with cyclic blasting excavation. When the damage depth at a certain location continuously increases or the damage variable reaches a high value area, that location is identified as a key control area. In this area, measures such as strengthening support, optimizing the delay section, adjusting the single-stage charge quantity, or shortening the single-cycle advance are implemented.

[0128] The above steps constitute a progressive active control process: grouting pretreatment cuts off the conditions for water-induced damage; SHPB dynamic testing and numerical inversion obtain the true dynamic constitutive parameters of the surrounding rock; numerical simulation based on calibration parameters determines the damage-sensitive area and optimizes the blasting design scheme; zonal calculation and application of equivalent blasting load differentiates the blasting energy according to borehole type; directional charging and detonation sequence setting control the input amplitude of damage from the source of blasting energy; pre-reinforcement enhances the initial resistance of the damage-sensitive area; dynamic control of blasting intensity and immediate support impose constraints at the final stage of damage generation and propagation; support parameters are determined based on numerical simulation results to match the bearing capacity with the surrounding rock pressure; and the collection and analysis of monitoring section data provides feedback for subsequent cycle parameter adjustments. Each step is implemented sequentially along the construction timeline. The results of the previous stage provide improved implementation conditions for the next stage, and the operational effect of the next stage is built on the favorable foundation established in the previous step, forming a closed loop of active control of surrounding rock damage throughout the entire process of cyclic blasting construction of small-clearance tunnels.

[0129] In alternative implementations of this scheme, besides the RHT model, the dynamic constitutive model can also employ the HJC model or other dynamic constitutive models for rocks that can consider strain rate effects and damage evolution. These alternative models, after parameter calibration through dynamic impact tests, can also output surrounding rock damage variables and be used for cyclic blasting numerical simulations. The equivalent blasting load can be obtained not only through theoretical calculations but also through inversion of field blasting vibration monitoring data, measured borehole pressure, or back analysis based on measured data. This equivalent blasting load time history can then be applied as input conditions to the corresponding excavation boundary of the numerical model. The evaluation of the surrounding rock damage state can be based on peak vibration velocity, damage variables, and damage depth, as well as a comprehensive evaluation combining surrounding rock stress, plastic zone range, dominant vibration frequency, particle acceleration, acoustic wave velocity attenuation, or field crack investigation results. The application scenarios of this scheme extend beyond small-clearance highway tunnels to include railway tunnels, urban underground tunnels, mine roadways, and other underground engineering projects constructed using the drill-and-blast method where adjacent chambers experience mutual disturbance.

[0130] The above provides a detailed description of the active control method for surrounding rock damage in small-clearance tunnel cyclic blasting. Specific examples are used to illustrate the principle and implementation of the invention. The description of the above embodiments is only for the purpose of helping to understand the method and core idea of ​​the invention. At the same time, for those skilled in the art, there will be changes in the specific implementation and application scope based on the idea of ​​the invention. Therefore, the content of this specification should not be construed as a limitation of the invention.

Claims

1. A method for actively controlling surrounding rock damage in cyclic blasting of tunnels with small clearance, characterized in that, The method includes: Grouting is used to fill the original fissures in the surrounding rock mass of the tunnel. For the surrounding rock after grouting, blast holes with continuous charge structure are set at the center of the tunnel face, and blast holes with interval charge structure are set at the tunnel excavation outline. The detonation sequence is set so that the blast holes at the center of the tunnel face are detonated before the blast holes at the excavation outline. The arch shoulder, arch foot, and interbedded rock areas of the surrounding rock are pre-reinforced; After pre-reinforcing the arch shoulder, arch foot, and interbedded rock area, blasting excavation is carried out. In each blasting excavation, the net distance between the advance tunnel and the follow tunnel, the surrounding rock grade, and the allowable peak vibration velocity are obtained. Based on the net distance, the surrounding rock grade, and the allowable peak vibration velocity, the maximum charge threshold for a single section is determined. The actual charge amount for a single section is set to not exceed the maximum charge threshold for a single section. After each blasting excavation is completed, a support structure is applied to the excavation face.

2. The method according to claim 1, characterized in that, The grouting filling uses cement grout with a water-cement ratio ranging from 0.6:1 to 1.2:1 and a grouting pressure ranging from 0.5MPa to 2.0MPa. The grouting holes are arranged in a quincunx pattern with a spacing of 1.0m to 2.0m. The grouting reinforcement depth extends no less than 3m beyond the tunnel excavation outline. The grouting is completed when the grouting pressure reaches the designed final pressure and the grouting volume is less than the preset threshold.

3. The method according to claim 2, characterized in that, When the blast hole at the center of the working face adopts a continuous charging structure, the ratio of the charge diameter to the blast hole diameter is within the range of the radially uncoupled charging coefficient. A concentrated charge pack is set at the bottom of the blast hole at the center of the working face. The concentrated charge pack is filled with a predetermined proportion of the designed charge within a predetermined length range at the bottom of the blast hole. When the blast hole at the excavation contour position adopts an intermittent charging structure, the intermittent material is air or an inert solid medium, and the length of a single intermittent segment is determined according to the blast hole depth and the number of charge segments.

4. The method according to claim 3, characterized in that, The detonation sequence is as follows: the slotted hole at the center of the working face is detonated first to create a free surface. After the slotted hole is detonated, the auxiliary holes, peripheral holes and bottom edge holes are detonated in sequence. The delay time between adjacent sections meets the conditions for staggered propagation of blasting vibration waves.

5. The method according to claim 4, characterized in that, The pre-reinforcement of the surrounding rock at the arch shoulder, arch foot, and interbedded rock area includes: installing advanced small guide pipes and grouting at the arch shoulder and arch foot, wherein the outer insertion angle of the advanced small guide pipes is determined according to the tunnel excavation outline, and the circumferential spacing is determined according to the degree of fracture of the surrounding rock; and installing tie rods in the interbedded rock area, wherein the tie rods penetrate the interbedded rock and apply prestress.

6. The method according to claim 5, characterized in that, The determination of the single-segment maximum charge threshold includes: during the initial blasting excavation, selecting a site coefficient and attenuation index from the empirical values ​​of blasting vibration attenuation parameters for the same level of surrounding rock as specified in the blasting safety regulations, based on the surrounding rock level; calculating the single-segment maximum charge threshold for the initial blast based on the site coefficient, the attenuation index, the allowable peak vibration velocity, and the distance from the blast source to the monitoring point; and during subsequent cyclic blasting excavation, fitting a corrected site coefficient and attenuation index based on the on-site blasting vibration monitoring data of each completed blast, and updating the single-segment maximum charge threshold based on the corrected site coefficient and attenuation index.

7. The method according to claim 6, characterized in that, The applied support structure includes: spraying a first concrete layer on the surface of the excavation face; erecting a steel arch frame on the surface of the first concrete layer along the tunnel excavation outline; drilling anchor bolt holes and installing system anchor bolts at the arch crown, arch shoulder, arch waist, and arch foot positions of the excavation face; spraying a second concrete layer to the designed thickness after the steel arch frame and system anchor bolts are installed; after the second concrete layer is sprayed, drilling locking anchor bolt holes and installing locking anchor bolts at the arch foot positions; the installation direction of the locking anchor bolts is inclined at an angle to the horizontal plane; the number of locking anchor bolts installed at the same arch foot position is determined according to the degree of fracture of the surrounding rock at the arch foot.

8. The method according to claim 1 or 7, characterized in that, Before grouting and filling the original fissures in the surrounding rock mass of the tunnel, the following steps are also included: sampling the surrounding rock of the tunnel and conducting SHPB dynamic impact tests to obtain the dynamic stress-strain curves and failure modes of the surrounding rock under different impact velocities, establishing a numerical model with the same size as the SHPB test device, and adjusting the parameters of the RHT dynamic constitutive model with the dynamic stress-strain curves and failure modes as calibration targets.

9. The method according to claim 8, characterized in that, After adjusting the parameters of the RHT dynamic constitutive model, the process also includes: establishing a numerical model for cyclic blasting excavation of a small-clearance tunnel based on the preliminary borehole design scheme; substituting the calibrated RHT dynamic constitutive model parameters into the numerical model; simulating multiple blasting excavation processes of the lead tunnel and the follow tunnel; extracting damage variable cloud maps and damage depth data; identifying locations where the damage variable exceeds a preset damage threshold and the damage depth exceeds a preset depth threshold as damage-sensitive areas requiring pre-reinforcement; and adjusting the charge structure and detonation sequence of the boreholes at the center of the tunnel face and the boreholes at the excavation contour locations based on the simulation results.

10. The method according to claim 9, characterized in that, The method further includes: obtaining the curves of the deformation of the surrounding rock and the stress on the support structure with the excavation cycle through numerical simulation of cyclic blasting; determining the construction parameters of the support structure based on the curves; and determining the specifications of the support structure so that the bearing capacity of the support structure is greater than the simulated surrounding rock pressure.

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