A method and system for screening blasting parameters in the one-time shaft sinking blasting of a blind raise

Abstract

The present invention belongs to the field of underground blasting technology, and discloses a method and system for screening blasting parameters in the one-time shaft formation blasting of blind raises. The method includes: determining the cut method, and further determining the cut form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length parameter; calculating the charge amount, and determining the reasonable charge amount and decoupling coefficient of the cut holes and the perimeter holes through numerical simulation calculation and the results of blasting funnel tests; determining the delay time, and calculating the optimal influence time of the hole-to-hole delay on the superposition of explosion stress waves by means of numerical simulation, so that the damage effect of the stress waves on the rock reaches the best; stemming and charging, because upward decoupled charging is adopted, determining a reasonable charging method to ensure the accuracy of the charge amount and the position of the explosive, and setting reasonable stemming to ensure that the explosive energy fully acts on the blasted rock mass. The present invention can greatly improve the safety of shaft formation operations; realize the maximization of resource utilization and improve economic benefits.

Description

Technical Field

[0001] The present invention belongs to the technical field of underground blasting, and particularly relates to a method and system for screening blasting parameters in the one-time shaft formation blasting of a blind raise. Background Technique

[0002] Regarding the underground blasting technology, domestic and foreign scholars have formed a scientific and complete system, and in-depth research has been carried out on rock property testing and different blasting parameter designs, and certain research results have been obtained. In terms of the calibration of rock mechanics parameters, Peng Jianyu carried out relevant experimental and numerical simulation studies on the fracture behavior of sandstone under Hopkinson bar impact load under static stress and the blasting fragmentation behavior of cement mortar specimens under static stress for the problem of the formation of blasting funnels under static stress. The phenomena and laws in the dynamic fracture process of rocks under static stress were studied, and the failure mechanism of rocks under the action of static and dynamic loads was revealed. In terms of blasting funnels, Wang Peng et al. used the ANSYS / LS-DYNA nonlinear three-dimensional dynamic finite element software to numerically simulate the stress distribution and propagation mechanism of rocks under the action of multi-hole simultaneous blasting. The stress distribution nephogram at different times and the stress-time history curve of typical elements were obtained. According to the results, the stress wave propagation law and the formation process of the blasting funnel were studied, and the relevant factors affecting the formation of the blasting funnel were discussed. And in terms of the classification of rock blastability, Xue Jianguang et al. established an attribute recognition model for the classification and discrimination of rock mass blastability in engineering blasting; the density, tensile strength, impact dynamic load strength, and rock mass integrity coefficient of the rock were selected as the discrimination indexes for attribute recognition, effectively solving the problem of evaluating the difficulty of rock mass blastability.

[0003] In the research on the key technologies of the one-time shaft formation blasting of the upward blind raise, the key in the construction of the one-time shaft formation technology of the upward fan-shaped cutting raise lies in the arrangement of the middle cut holes and the peripheral smooth holes. A reasonable center cut hole diameter and hole layout method are selected to ensure the cut quality, and blasting parameters such as the decoupling coefficient, hole spacing, and delay time of the peripheral holes are calculated to ensure the size and quality of the cutting raise to meet the requirements of subsequent construction.

[0004] Through the above analysis, the problems and defects of the existing technology are as follows: In the construction of the one-time shaft formation technology of the existing upward fan-shaped cutting raise, the control effect of relevant parameters such as the number of expansion holes, delay time, and charge amount is poor, and it cannot ensure that the shaft size and the stability of the surrounding reserved rock mass meet the design requirements. Summary of the Invention

[0005] To overcome the problems existing in the related technology, the disclosed embodiments of the present invention provide a method and system for screening blasting parameters in the one-time shaft formation blasting of a blind raise.

[0006] The technical solution is as follows: A method for screening blasting parameters in the one-time shaft formation blasting of a blind raise, including the steps:

[0007] S1. Determine the cut - hole method. According to the mining operation equipment and the results of numerical simulation calculation and analysis, select a cut - hole scheme suitable for lithium ore rock and conduct on - site tests to determine the parameters of the cut - hole form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length.

[0008] S2. Calculate the charge amount. Based on the results of numerical simulation calculation and blasting funnel tests, determine the charge amount and decoupling coefficient of the cut - hole and perimeter holes.

[0009] S3. Determine the delay time. Use numerical simulation to calculate the influence time of the delay between holes on the superposition of explosion stress waves to achieve the destruction of rocks by stress waves.

[0010] S4. Stemming and charging. Determine the charging method to ensure the accuracy of the charge amount and the position of the explosive, and set reasonable stemming to ensure that the explosive energy fully acts on the blasted rock mass.

[0011] In step S2, the decoupling coefficient is the ratio of the hole diameter to the charge diameter; the value of the decoupling coefficient is 1.5 - 3.0; the perimeter hole spacing is 10 - 20 times the hole diameter; the hole density coefficient of the perimeter holes is taken as 0.8 - 1.0; the charge concentration of the perimeter holes is 70 - 120 kg / m in soft rock, 100 - 150 kg / m in medium - hard rock, and 150 - 250 kg / m in hard rock.

[0012] Furthermore, the method for determining the decoupling coefficient includes:

[0013] According to the initially determined ratio of the decoupled hole diameter to the charge diameter, preset the decoupling coefficient for the rock mass. After determining the decoupled charge explosive material model, using the established JWL equation of state for explosion action and the LINEAR - POLYNOMIAL equation of state for air fluid, calculate whether the tangential tensile stress generated by the decoupled charge explosive material models with different decoupling coefficients is greater than the tensile strength of the rock, and determine the ratio of the decoupled hole diameter to the charge diameter of the rock mass and the decoupling coefficient that meet the energy and resolution as the parameters for generating directional fractures by the optimized decoupled charge explosive material model.

[0014] Furthermore, calculating whether the tangential tensile stress generated by the decoupled charge explosive material models with different decoupling coefficients is greater than the tensile strength of the rock includes:

[0015] (1) During the explosion process, simplify the explosive material model into an incompressible Newtonian explosive fluid. According to Newton's second law, obtain the motion differential equation of the explosive fluid; based on the continuity equation derived from the mass conservation equation in the volume element, obtain the formula for the tangential tensile stress generated instantaneously by the explosive fluid, and calculate the distribution of the tangential tensile stress of the fluid in the cut - hole, which should couple the flow field and the pressure field.

[0016] (2) After the explosion of the explosive material model, the tangential tensile stress generated during the conversion into air fluid is determined by the tangential tensile stress conduction equation;

[0017] (3) During the generation process of explosion cracks, the rock mass is simplified as a rigid deformation model; before the rock mass reaches the fracture strength, the tangential tensile stress and strain generated by the explosion of the explosive material model show a linear relationship. After the stress reaches the fracture strength of the rock mass, it remains constant, and the final strain is the sum of the fracture strain and the crack strain.

[0018] In step (1), according to Newton's second law, the motion differential equation of the explosive fluid is obtained, and the expression is:

[0019]

[0020] In the formula, ρ' is the density of the explosive fluid, o, c, f are the pressure motion velocity components of the explosive fluid in the x, y, z axes respectively, e is the explosion motion time of the explosive fluid, h is the pressure per unit volume of the explosive fluid, kx, ky, kz are the explosion pressure acceleration components in the three coordinate axis directions respectively, δ is the density of the explosive fluid after explosion motion, is the Laplace operator.

[0021] In step (1), according to the continuity equation derived from the mass conservation equation in the volume element, the formula for the instantaneous tangential tensile stress of the explosive fluid is obtained, and the expression is:

[0022]

[0023] In the formula, I is the dispersion degree of the explosive fluid;

[0024] To calculate the distribution of the tangential tensile stress of the fluid in the groove hole, the flow field and the pressure field should be coupled, and the energy equation is as follows:

[0025]

[0026] In the formula, is the slope of the tangential tensile stress - pressure curve of the explosive material model, A is the tangential tensile stress, and J is the pressure conductivity of the explosive material model.

[0027] In step (2), after the explosion of the explosive material model, the tangential tensile stress generated during the conversion into air fluid is determined by the tangential tensile stress conduction equation, and the expression is:

[0028]

[0029] In the formula, r is the internal energy of the explosive material model.

[0030] In step (3), it should finally become the sum of the fracture strain and the crack strain, and the expression is:

[0031]

[0032] In the formula, is the theoretical stress of the rock mass, E is the fracture modulus, is the theoretical strain of the rock mass, and η s is the stress when cracks are generated in the rock mass, is the strength when the rock mass fractures, are the strain corresponding to the fracture of the rock mass and the crack strain generated subsequently, respectively.

[0033] In step S3, the delay time is determined, including the steps:

[0034] S301. Through the three-dimensional lithium ore rock structure diagram, determine the axial height difference D and the straight-line distance l between the pre-blast area of the ore rock and the blast center of the blast holes in the blast area; measure the shear wave velocity u of the blasting vibration wave in the rock mass h and the longitudinal wave velocity u z ; determine the structural influence coefficient G2 and the axial height influence coefficient χ of the blast holes through blasting vibration monitoring;

[0035] In the same accurate delay controlled blasting, the damping ratio of the vibration is the same. The blasting vibration velocity is affected by the axial height condition of the blast holes. The blasting vibration velocity is expressed as:

[0036] U (s) = sinθ × G2(D / l) χ

[0037] In the formula, U (s) is the blasting vibration velocity, and θ is the included angle formed by the axial height difference D and the straight-line distance l of the blast holes;

[0038] S302. According to the blasting vibration velocity calculation formula, the centroid vibration frequency of the lithium ore rock mass conforms to the dimensional analysis theorem and the π theorem. According to the straight-line distance between the pre-blast area of the ore rock and the blast center of the blast holes and the axial height difference, calculate the centroid vibration frequency q of the lithium ore rock mass as:

[0039] q = F · G2(D / l) χ

[0040] In the formula, F is the frequency coefficient;

[0041] S303. By setting the inter-hole delay time, make the adjacent vibration waveforms differ by N / 4 cycles when reaching the specified blast hole node. N is an integer, so that the wave peaks of the two waveforms meet when reaching this position, so that the amplitudes are superimposed. The inter-hole delay time is calculated according to the following formula:

[0042] According to the calculation formula for the centroid vibration frequency of lithium ore rock mass, let the distances between the pre-blast area of the ore rock and the blast holes be \(l_1, l_2, \cdots, l\) n , \(l_1 \lt l_2 \lt \cdots \lt l\) n , and the axial height differences of the blast holes be \(D_1, D_2, \cdots, D\) n , and the initiation times of the blast holes be \(s_1, s_2, \cdots, s\) n , \(s_1 \lt s_2 \lt \cdots \lt s\) n ; meanwhile, the difference in the distances between the blast holes and the pre-blast area of the ore rock is \(\Delta l = l\) n - l n-1 , and the interval time between holes is \(\Delta s = s\) n - s n-1 ; the expression is:

[0043]

[0044] In the formula, \(l\) n is the distance between the pre-blast area of the ore rock and the \(n\)th blast hole, and \(l\) n-1 is the distance between the pre-blast area of the ore rock and the \((n - 1)\)th blast hole, and \(s\) n is the initiation time of the \(n\)th blast hole, and \(s\) n-1 is the initiation time of the \((n - 1)\)th blast hole.

[0045] Another object of the present invention is to provide a blasting parameter screening system for one-time well completion blasting of blind raises. This system implements the method for screening blasting parameters in the one-time well completion blasting of blind raises, and the system includes:

[0046] A cut pattern determination module, which is used to select a cut pattern suitable for lithium ore rock according to the mine operation equipment and the results of numerical simulation calculation and analysis, and conduct on-site tests to determine the parameters of the cut form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length;

[0047] An uncoupling coefficient determination module, which is used to determine the charge amount and uncoupling coefficient of the cut holes and the perimeter holes through numerical simulation calculation and the results of blasting funnel tests;

[0048] An inter-hole delay determination module, which is used to calculate the influence time of the inter-hole delay on the superposition of explosion stress waves by numerical simulation to achieve the destruction of the rock by the stress waves;

[0049] A stemming and charging module, which is used to determine the charging method to ensure the accuracy of the charge amount and the position of the explosive, and set a reasonable stemming to ensure that the explosive energy fully acts on the rock mass to be blasted.

[0050] Combining all the above technical solutions, the beneficial effects of the present invention are as follows: The design of the blasting parameters for the one-time well completion of the blind raise in the present invention can realize the uphole blind raise one-time well completion technology, which can greatly improve the safety of the well completion operation; realize the maximization of resource utilization and improve economic benefits. The present invention calculates the empty hole size and quantity in the uphole blind raise one-time well completion blasting technology to ensure sufficient compensation space during blasting. Strictly control relevant parameters such as the number of reaming holes, delay time, and charge amount to ensure that the well completion size and the stability of the surrounding reserved rock mass meet the design requirements. BRIEF DESCRIPTION OF THE DRAWINGS

[0051] The accompanying drawings herein are incorporated into the specification and constitute a part of this specification, showing embodiments consistent with the present disclosure, and are used together with the specification to explain the principles of the present disclosure;

[0052] Figure 1 is a flowchart of a method for screening blasting parameters in the uphole blind raise one-time well completion blasting provided by an embodiment of the present invention;

[0053] Figure 2 is a schematic diagram of a system for screening blasting parameters in the uphole blind raise one-time well completion blasting provided by an embodiment of the present invention;

[0054] In the figure: 1, cut method determination module; 2, decoupling coefficient determination module; 3, hole-to-hole delay determination module; 4, stemming and charging module. DETAILED DESCRIPTION OF THE EMBODIMENTS

[0055] To make the above objects, features, and advantages of the present invention more apparent and understandable, the following detailed description of the specific embodiments of the present invention is made in conjunction with the accompanying drawings. Many specific details are set forth in the following description to fully understand the present invention. However, the present invention can be implemented in many other ways different from those described herein, and those skilled in the art can make similar improvements without departing from the connotation of the present invention. Therefore, the present invention is not limited by the specific embodiments disclosed below.

[0056] Example 1, as Figure 1 shown, the method for screening blasting parameters in the uphole blind raise one-time well completion blasting provided by an embodiment of the present invention includes:

[0057] S1. Determine the cut method. According to the mining operation equipment and the results of numerical simulation calculation and analysis, select a cut scheme suitable for the lithium ore rock and conduct on-site tests to determine the parameters of the cut form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length;

[0058] S2. Calculate the charge amount. Through the results of numerical simulation calculation and blasting funnel test, determine the charge amount and decoupling coefficient of the cut holes and the perimeter holes;

[0059] S3. Determine the delay time, and calculate the influence time of the delay between holes on the superposition of explosion stress waves by numerical simulation to achieve the destruction of rocks by stress waves.

[0060] S4. Tamping and charging. Determine the charging method to ensure the accuracy of the charge amount and the position of the explosive, and set reasonable tamping to ensure that the energy of the explosive fully acts on the rock mass to be blasted.

[0061] Exemplarily, in step S2, the decoupling coefficient in the blasting parameters refers to the ratio of the hole diameter to the charge diameter. In controlled blasting, the decoupling coefficient is a very important parameter, mainly used for presplitting blasting and smooth blasting. Its purpose is to protect the integrity of the blasting, prevent cracking and reduce fissures, and maintain the stability of the rock mass. The value of the decoupling coefficient generally ranges from 1.5 to 3.0. At this time, the impact pressure (or the generated stress) on the rock on the hole wall can be made not greater than the ultimate compressive strength of the rock, avoiding the formation of a crushed zone, thus leaving a semi-hole trace. At the same time, the tangential tensile stress generated on the hole center line is greater than the tensile strength of the rock, generating directional fissures.

[0062] During smooth blasting, when the decoupling coefficient of the perimeter holes is taken as 2 - 5, the smooth blasting effect is the best; the spacing of the perimeter holes is generally taken as the hole distance is 10 - 20 times the hole diameter; the hole density coefficient of the perimeter holes is generally taken as 0.8 - 1.0; the charge concentration of the perimeter holes is generally 70 - 120 kg / m in soft rock, 100 - 150 kg / m in medium-hard rock, and 150 - 250 kg / m in hard rock.

[0063] Exemplarily, the method for determining the decoupling coefficient includes:

[0064] According to the ratio of the initially determined decoupled hole diameter to the charge diameter, a series of corresponding decoupling coefficients are preset for the rock mass. After the decoupled charge explosive material model is determined, use the established JWL equation of state for explosion action and the LINEAR - POLYNOMIAL equation of state of air fluid to calculate whether the tangential tensile stress generated by the decoupled charge explosive material model with different decoupling coefficients is greater than the tensile strength of the rock, and determine the ratio of the decoupled hole diameter to the charge diameter of the rock mass and the decoupling coefficient that meet the energy and resolution as the parameters for generating directional fissures by the optimized decoupled charge explosive material model.

[0065] Exemplarily, using the established JWL equation of state for explosion action and the LINEAR - POLYNOMIAL equation of state of air fluid to calculate whether the tangential tensile stress generated by the decoupled charge explosive material model with different decoupling coefficients is greater than the tensile strength of the rock includes:

[0066] (1) During the explosion process, the explosive material model is simplified to an incompressible Newtonian explosive fluid. According to Newton's second law, the differential equation of motion of the explosive fluid is obtained:

[0067]

[0068] Where p' is the density of the explosive fluid, o, c, f are the pressure motion velocity components of the explosive fluid in the x, y, z axes respectively, e is the explosion motion time of the explosive fluid, h is the pressure per unit volume of the explosive fluid, kx, ky, kz are the explosion pressure acceleration components in the three coordinate axis directions respectively, δ is the density of the explosive fluid after the explosion motion, is the Laplace operator.

[0069] According to the continuity equation derived from the mass conservation equation in the volume element, the formula for the tangential tensile stress generated instantaneously by the explosive fluid is obtained:

[0070]

[0071] Where I is the dispersion degree of the explosive fluid;

[0072] To calculate the tangential tensile stress distribution of the fluid in the shaped groove hole, the flow field and the pressure field should be coupled. The energy equation is as follows:

[0073]

[0074] Where is the slope of the tangential tensile stress-pressure curve of the explosive material model, A is the tangential tensile stress, and J is the pressure conductivity of the explosive material model.

[0075] In step (2), after the explosion of the explosive material model, the tangential tensile stress generated during the conversion to the air fluid is determined by the tangential tensile stress conduction equation, and the expression is:

[0076]

[0077] Where r is the internal energy of the explosive material model.

[0078] In step (3), the final strain is the sum of the fracture strain and the crack strain, and the expression is:

[0079]

[0080] Where is the theoretical stress of the rock mass, E is the fracture modulus, is the theoretical strain of the rock mass, η s is the stress when cracks are generated in the rock mass, is the strength when the rock mass fractures, are the corresponding strain and the subsequent crack strain when the rock mass fractures respectively.

[0081] Exemplarily, in step S3, the determination of the delay time includes:

[0082] S301. Determine the axial height difference D and the straight-line distance l between the pre-blast area of the ore-rock and the blast center of the blast holes through the three-dimensional lithium ore-rock structure diagram; measure the shear wave velocity u of the blasting vibration wave in the rock mass h and the longitudinal wave velocity u z ; determine the structural influence coefficient G2 and the axial height influence coefficient χ of the blast holes through blasting vibration monitoring;

[0083] Assume that the damping ratio of the vibration is the same in the same precise delay-controlled blasting. The blasting vibration velocity is affected by the axial height condition of the blast holes, and the blasting vibration velocity is expressed as:

[0084] U (s) = sinθ × G2(D / l) χ

[0085] In the formula, U (s) is the blasting vibration velocity, θ is the included angle formed by the axial height difference D and the straight-line distance l;

[0086] S302. According to the blasting vibration velocity calculation formula, the centroid vibration frequency of the lithium ore-rock mass conforms to the dimensional analysis theorem and the π theorem. According to the straight-line distance and the axial height difference between the pre-blast area of the ore-rock and the blast center of the blast holes, calculate the centroid vibration frequency q of the lithium ore-rock mass as:

[0087] q = F · G2(D / l) χ

[0088] In the formula, F is the frequency coefficient;

[0089] S303. By setting the delay time between holes, make the adjacent vibration waveforms differ by N / 4 cycles when reaching the specified blast hole node, N is an integer, so that the wave peaks of the two waveforms meet when reaching this position, and thus the amplitudes are superimposed. The delay time between holes is calculated as follows:

[0090] According to the centroid vibration frequency calculation formula of the lithium ore-rock mass, let the distances between the pre-blast area of the ore-rock and the blast holes be l1, l2…l n , l1 < l2 < … < l n , the axial height differences of the blast holes are D1, D2…D n , the initiation times of the blast holes are s1, s2…s n , s1 < s2 < … < s n ; at the same time, the difference in the distances between the blast holes and the pre-blast area of the ore-rock is Δl = l n -l n-1 , the interval time between holes is Δs = s n -sn-1 ; The expression is:

[0091]

[0092] In the formula, l n is the distance between the pre-blast area of ore and rock and the nth blast hole, and l n-1 is the distance between the pre-blast area of ore and rock and the (n - 1)th blast hole, s n is the detonation time of the nth blast hole, in seconds, and s n-1 is the detonation time of the (n - 1)th blast hole.

[0093] Coupled charge: In this structure, the explosive does not completely fill the blast hole, but there is a certain gap between the explosive and the hole wall. Air-decked charge is a typical uncoupled charge. It uses air as the spacer medium, which can increase the height of the charge column, reduce the amount of explosive charge, prolong the action time of the explosive gas in the blast hole, increase the action area of the explosive gas, and reduce the damage of the blast shock wave to the hole wall.

[0094] Exemplarily, there are two charging methods: forward charging and reverse charging. The primer is located at the hole mouth, the shaped charge cavity faces the bottom of the hole, and the detonation direction is from the hole mouth to the bottom of the hole. This kind of charging is forward charging, and the blasting operation with forward charging is forward blasting. Reverse charging is the opposite.

[0095] In blasting operations, reverse blasting is superior to forward blasting; it can make full use of the energy of the explosive, and the movement time of the stemming is relatively late.

[0096] The blasting efficiency is high. The detonator cartridge is located at the bottom of the hole, which can ensure that the explosive fully reacts, with a large explosion energy, and is conducive to overcoming the clamping effect of the rock at the bottom of the hole. Due to the channel effect in forward blasting, the farther away from the detonator cartridge, the lower the detonation velocity and the smaller the blasting effect, which is not conducive to overcoming the clamping effect of the rock at the bottom of the hole. If a misfire occurs, forward blasting is prone to produce a dud, and the primer can explode outside the blast hole; while reverse blasting will not throw the detonator out of the blast hole. Judging from the flame generated by blasting, under the condition of not installing stemming, the flame generated by reverse blasting is longer than that of forward blasting. Considering the actual level and quality of underground workers, when millisecond blasting is used in the excavation and working faces of high-gas mines and high-gas areas of low-gas mines, if reverse detonation is adopted, safety technical measures must be formulated.

[0097] In setting reasonable stemming, it includes:

[0098] Continuous charge: The explosive is continuously loaded along the axial direction of the blast hole. When the hole depth exceeds 8 m, generally two detonator cartridges are arranged, one is placed 0.3 - 0.5 m away from the bottom of the hole, and the other is placed 0.5 m at the top of the charge column. The advantage is simple operation; the disadvantage is that the charge column is relatively low, and large lumps are prone to occur in the uncharged part at the hole mouth.

[0099] Sectional charging: The charge column in the deep hole is divided into several sections, separated by air, rock debris or water. The advantages are that it increases the charging height and reduces the generation of large pieces at the orifice part; the disadvantage is that the construction is troublesome.

[0100] Bottom interval charging: Leave a section of length at the bottom of the deep hole without charging, with air as the interval medium; in addition, there are also water intervals and flexible material intervals. Implementing air interval charging at the bottom of the hole is also called bottom air cushion charging.

[0101] Mixed charging: A mixed charging method where high-power explosives are loaded at the bottom of the hole and ordinary explosives are loaded in the upper part.

[0102] Example 2, to further describe the technical features related to the present invention, the content is as follows:

[0103] (1) Static and dynamic tests of lithium ore rock mass and its explosibility classification.

[0104] (1.1) Static tests and blasting vibration wave tests.

[0105] Conduct static mechanical tests on ore rocks with different lithologies and different degrees of joint development to obtain mechanical parameters such as static tensile and compressive strengths, and conduct blasting vibration wave tests to obtain basic parameters such as wave velocity, so as to provide data support for subsequent tests.

[0106] (1.2) Dynamic mechanical property tests of rock mass.

[0107] a. Conduct dynamic mechanical property tests on ore rocks. Through the SHPB dynamic impact test device, conduct dynamic compression and dynamic tensile tests to obtain the dynamic crushing energy consumption curve of ore rocks, and clarify the influence of the degree of joint development of ore rocks on rock fragmentation and energy consumption.

[0108] b. Conduct dynamic mechanical property tests on rock mass with confining pressure. Conduct dynamic mechanical tests under uniaxial and triaxial confining pressure conditions to obtain the stress changes and rock mass deformation characteristics during the dynamic failure process of rock mass under in-situ stress conditions, obtain the deformation characteristics and time curve of rock mass confining pressure during dynamic excavation, and clarify the influence of dynamic changes in confining pressure on the mechanical properties of lithium ore rock.

[0109] (1.3) Explosibility classification.

[0110] By sorting out and analyzing many parameters such as rock compressive strength, bulk density, rock integrity and engineering geological parameters, establish a weight model through mathematical analysis software, determine the key parameters affecting rock explosibility, and establish a simple and effective lithium ore explosibility classification standard according to the actual situation of the mine site, providing a standardized reference and systematic design basis for subsequent blasting parameter design.

[0111] (2) Blasting funnel test.

[0112] (2.1) Design blasting funnel tests according to the dynamic mechanical properties of ore and rock to determine the optimal specific charge for different types of surrounding rock and ore, and the influence laws of different joint development degrees on rock fragmentation and fragment throwing.

[0113] (2.2) Based on the results of single-hole blasting funnel tests, design multi-hole blasting funnel tests and conduct multi-hole cut blasting tests to determine the influence of parameters such as hole spacing and delay time on important evaluation parameters such as the volume and depth of the blasting funnel.

[0114] (3) Optimization of roadway tunneling blasting parameters.

[0115] (3.1) Optimization of cut hole blasting parameters.

[0116] (3.1.1) Research on the optimization of cut hole blasting parameters for jointed rock masses. For the jointed rock masses existing in roadway tunneling construction, conduct cut hole blasting tests. Through numerical simulation calculations and on-site tests, determine important parameters such as reasonable charge amount, decoupling coefficient, number of empty holes, hole spacing, and delay time in cut blasting of jointed rock masses, improve the cut quality, and provide sufficient additional space for subsequent blasting.

[0117] (3.1.2) Research on the optimization of cut hole blasting parameters for viscous rock masses of lithium ore. For the problems such as high large block rate and low rock fragmentation degree existing in the blasting process of lithium ore rock masses in roadway tunneling construction, conduct cut hole blasting tests. Through numerical simulation calculations and on-site tests, determine important parameters such as reasonable charge amount, number of empty holes, hole spacing, and delay time in cut blasting of viscous rock masses, improve the cut quality, and provide sufficient additional space for subsequent blasting.

[0118] (3.2) Optimization of perimeter smooth blasting parameters.

[0119] (3.2.1) Research on the optimization of perimeter smooth blasting parameters for jointed rock masses. For the jointed rock masses existing in roadway tunneling construction, conduct smooth blasting tests. Through numerical simulation calculations and on-site tests, determine important parameters such as reasonable charge amount, decoupling coefficient, hole spacing, and delay time in smooth blasting of jointed rock masses, improve the half-hole rate, footage length, etc., ensure the integrity of the roadway surrounding rock, and improve the safety and stability of the surrounding rock.

[0120] (3.2.2) Research on the optimization of perimeter smooth blasting parameters for viscous rock masses of lithium ore. For the problems such as low explosibility and low rock fragmentation degree existing in the blasting process of lithium ore rock masses in roadway tunneling construction, conduct smooth blasting tests. Through numerical simulation calculations and on-site tests, determine important parameters such as reasonable charge amount, decoupling coefficient, hole spacing, and delay time in smooth blasting of viscous rock masses, improve the half-hole rate, footage length, etc., ensure the integrity of the roadway surrounding rock, and improve the safety and stability of the surrounding rock.

[0121] (3.3) Research on the full-face blasting design for roadway excavation;

[0122] (3.3.1) Research on the optimization of blasting parameters for roadway excavation in joint-developed rock mass. Combining the blasting funnel test, cut blasting test, and smooth blasting test, formulate a reasonable blasting network design for roadway excavation. Through numerical simulation and on-site tests, select the optimal combination of parameters such as charge amount, hole spacing, delay time, and decoupling coefficient to ensure that important indicators such as roadway footage and roadway boundary meet the relevant design requirements during the roadway excavation construction in joint-developed rock mass.

[0123] (3.3.2) Research on the optimization of blasting parameters for roadway excavation in viscous rock mass of lithium ore. Combining the blasting funnel test, cut blasting test, and smooth blasting test, formulate a reasonable blasting network design for roadway excavation. Through numerical simulation and on-site tests, select the optimal combination of parameters such as charge amount, hole spacing, delay time, and decoupling coefficient to ensure that important indicators such as roadway footage and roadway boundary meet the relevant design requirements during the roadway excavation construction in viscous rock mass of lithium ore.

[0124] (4) Research on the key technologies of upward blind raise blasting in one pass.

[0125] The key in the construction of the upward fan-shaped cut raise blasting in one pass lies in the arrangement of the middle cut holes and the peripheral smooth holes. Select a reasonable center cut hole diameter and hole layout method to ensure the cut quality, and calculate blasting parameters such as the decoupling coefficient, hole spacing, and delay time of the peripheral holes to ensure the size and quality of the cut raise to meet the requirements of subsequent construction.

[0126] 1) Research on cut blasting parameters.

[0127] a. Determination of the range of the explosion stress wave fracture zone. Cut blasting is carried out under the condition of only one free face, and the rock breaking is difficult. In the present invention, its rock breaking effect can be regarded as the superposition of the blasting effects of each cut hole. The charges of the simultaneously detonated cut holes form radial and circumferential cracks around the blast holes and intersect into a spatial crack network in the cut cavity, cutting the rock into fragments and throwing them out under the expansion action of the detonation gas. Therefore, to ensure sufficient rock fragmentation, the layout diameter of the cut holes must meet the design requirements.

[0128] b. Deep hole straight cut uses the space of the empty hole to provide a free face and compensation space for the rock blasted down. Therefore, the reserved space for each cut blasting should meet the swelling requirements of the rock. Determine the blasting compensation space requirements of the ore rock through numerical simulation and on-site tests to ensure the rock fragmentation effect. Determine the number and hole layout method of the middle empty holes through on-site tests.

[0129] 2) Research on blind raise blasting parameters.

[0130] After the project is launched, 3-5 optimization plans are comprehensively determined based on the rock mechanics parameters of the mine, the results of numerical simulation, and the calculation results of the blasting stress wave fracture circle determination method and the compensation space method, and on-site test research is carried out.

[0131] a. Determination of the cut method. At present, the common cut hole layout methods in the mine include nine-hole cut, rhombus cut, barrel cut, etc. According to the existing operation equipment in the mine and the results of numerical simulation calculation and analysis, the best cut plan suitable for the lithium ore rock is selected and on-site tests are carried out, and then parameters such as the cut form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length are determined.

[0132] b. Calculation of the charge amount. Through numerical simulation calculation and the results of the blasting funnel test, the reasonable charge amount and decoupling coefficient of the cut holes and the perimeter holes are determined.

[0133] c. Determination of the delay time. The numerical simulation method is used to calculate the best influence time of the hole-to-hole delay on the superposition effect of the explosion stress wave, so that the damage effect of the stress wave on the rock reaches the best.

[0134] d. Stemming and charging. Since upward decoupled charging is adopted, it is necessary to study a reasonable charging method to ensure the accuracy of the charge amount and the position of the explosive, and set reasonable stemming to ensure that the explosive energy fully acts on the blasted rock mass.

[0135] Example 3. The related technologies involved in the present invention also include:

[0136] (1) Static and dynamic mechanical property tests of rocks.

[0137] Analysis of the basic mechanical properties of rocks: In order to obtain the basic mechanical parameters of the rocks in the research area of the underground mine, a YAW-600 type pressure testing machine and other equipment are used for the static mechanical experiments of the rocks to measure mechanical parameters such as the uniaxial compressive strength, conventional triaxial compressive strength, shear resistance, tensile strength, and elastic modulus of the test pieces. The experimental process is operated according to the test methods recommended in the "Rock Test Regulations for Railway Engineering" (TB10115-2014). The longitudinal wave velocity of the rocks is measured by using an RSM-SY5 type blasting vibration wave tester and its supporting longitudinal wave transducer.

[0138] In this test, different axial pressures and confining pressures need to be applied to the rock test pieces, and equipment with an active confining pressure loading function is required. Therefore, a three-dimensional SHPB test system of a certain university is finally selected, which can apply an active confining pressure in the range of 0-100 MPa and an axial static pressure in the range of 0-200 MPa to the rock test pieces, can generate an impact load of 0-500 MPa, and can achieve loading in a large range of high strain rates (100-103 s-1). The test is completed in the rock mechanics laboratory of a certain university.

[0139] (2) Classification of the explosibility of ore rocks.

[0140] The explosibility of rock refers to the resistance of rock to blasting action or the ease of blasting rock, which is a comprehensive manifestation of the physical and mechanical properties of rock under dynamic load. The classification of explosibility is to divide rocks into grades of blasting difficulty according to the quantitative index of rock explosibility. It is an important basis for formulating blasting quotas, selecting blasting parameters, and conducting blasting design, and is also one of the scientific bases for the management of mining enterprises. Engineering practice has proved that the establishment of a reasonable classification standard has a significant effect on improving the quality of blasting construction, accelerating the project progress, and reducing the construction cost. Therefore, it is of great significance to accurately classify the explosibility of rock masses. The commonly used quantification methods are as follows:

[0141] 1) The Prokopyev rock strength classification.

[0142] The Prokopyev classification method. The rock strength coefficient f represents the relative value of the rock's resistance to fragmentation. Since the compressive strength of rock is the strongest, 1 / 10 of the uniaxial compressive strength limit of the rock is taken as the rock strength coefficient. The calculation formula of the rock strength coefficient is simple and clear, and the f value can be used to predict the rock's resistance to fragmentation and its stability after drilling. According to the rock strength coefficient f, rocks can be divided into 10 grades, and the higher the grade, the easier the rock is to break.

[0143] 2) The single-factor classification method of explosibility.

[0144] a. Classification by wave impedance.

[0145] The wave impedance of rock is the product of the longitudinal wave velocity and the rock density. It means the unit movement velocity generated by the centroid of the rock mass during blasting, and the magnitude of the stress that can be derived in the rock. A.H. Khanukayev's research on using the wave impedance of rock as the basis for blasting classification is a major progress in the study of rock blasting, because this index is measured in the in-situ rock mass, and the testing instruments and testing methods are relatively simple. A large number of experimental studies have shown that the wave impedance of the rock mass is not only related to the physical and mechanical properties of the rock, but also depends on the fracture structure characteristics of the rock.

[0146] b. Classification considering the blasting fragment size.

[0147] Different specific charges of explosives will result in different large block ratios. Rubtsov stipulated the following standard blasting conditions: the hole diameter is not greater than 0.02 times the bench height, the number of single-row simultaneous blasting holes is not less than 5, the overbreak is not greater than 0.15 times the burden, the hole spacing factor is 1, 6# waterproof ammonium nitrate explosive is used, continuous charging, the stemming coefficient is 0.5, and instantaneous initiation is adopted.

[0148] c. Classification by the standard explosive consumption q.

[0149] Rzhevsky suggested that the explosibility of rock be determined by the standard explosive consumption q, and q is closely related to fractures.

[0150] d. Optimal blasting funnel index.

[0151] When C.W. Livingston studied the law of loose blasting funnels, he developed a method to determine the explosibility of rocks through blasting funnel experiments. Livingston adopted the relationship that the minimum burden is proportional to the cube root of the explosive charge.

[0152] 3) Multi-factor classification method for explosibility.

[0153] a. Comprehensive explosibility classification.

[0154] This classification method comprehensively considers various factors such as the specific charge of explosives, rock hardness, and rock mass fractures, with the specific charge of explosives as the main factor. The standard conditions for the specific charge of explosives are: bench height 10 - 15m, hole diameter 243mm. Ammonium nitrate explosive, detonation heat 4190 KJ / Kg. A large amount of statistical data shows that the deviation (standard deviation) of the specific charge of explosives is proportional to the 2 / 3 power of the specific charge of explosives.

[0155] b. Comprehensive classification of rock explosibility.

[0156] This classification method mainly considers the relationship between the volume of the blasting funnel, the distribution of blasted fragment sizes, and the rock wave impedance and rock explosibility. The standard conditions are as follows: Directly select a representative rock section at the blasting site of the classified mine. On a relatively intact rock mass with one free face, drill vertical holes with a hole diameter of 45mm, a hole depth of 1m, and a hole spacing of 2m; Use No. 2 rock ammonium nitrate explosive, with a charge of 0.45 Kg per hole, continuous charging, stemming with stemming clay, and initiate with 1 No. 8 detonator. Test method: Use a blasting vibration wave instrument to measure the elastic longitudinal wave velocity of the rock mass before charging. After charging and blasting, measure the large fragment rate (greater than 300mm), small fragment rate (less than 50mm), and average qualification rate (cumulative average value of fragment sizes of 50 - 100mm, 100 - 200mm, and 200 - 300mm) of the muck rock, and measure and calculate the volume of the blasting funnel.

[0157] 4) Classification of rock mass explosibility using the grey system theory.

[0158] a. Selection of classification indicators.

[0159] There are dozens of factors affecting the quality of rock blasting, and their combined effects determine the quality of rock blasting. However, due to different purposes of theoretical analysis and field tests and various objective conditions, it is impossible and unnecessary to reflect all the influencing factors. Therefore, how to simply and reliably determine the main factors affecting the quality of rock blasting from numerous factors has become the basic problem in the theoretical and applied research of rock blasting and is also the prerequisite for accurately controlling and predicting the blasting quality. In an actual engineering system, often only part of the attributes or properties of the system are known, while the other part is unknown or uncertain. Therefore, it can be considered that the rock engineering geological system is a grey system. When it is required to evaluate the blastability grade of the system, according to the grey system theory, this is actually a hierarchical decision-making clustering problem. At this time, the system is described by grey parameters, that is, the grading indicators of the blastability grade are represented by grey numbers. In this way, the grey clustering method in the grey system theory can be used to evaluate the blastability grading of rock masses.

[0160] Following the following two principles: (1) It can comprehensively reflect the blastable attributes of rock masses from different aspects; (2) It can be relatively easily obtained through experimental or field test methods. The rock toughness coefficient f, the wave impedance of the rock, the unit explosive consumption, and the average fissure spacing of the rock mass can be used as the evaluation indicators of rock blastability.

[0161] b. Grey clustering grading of rock blastability.

[0162] Let k = 1, 2, 3, 4, 5 be the typical categories, i = Ⅰ, Ⅱ, Ⅲ, Ⅳ be the clustering elements, and j = 1#, 2#, 3#, 4# be the clustering indicators. The grey clustering grading method is to distinguish the category to which the clustering elements belong under the clustering indicators.

[0163] First, according to the research results and habits of rock blastability grading, the rocks are divided into five typical categories according to the ease of blastability, namely, easy to blast, medium, difficult to blast, very difficult to blast, and extremely difficult to blast, and they are regarded as the typical categories k according to the grey system theory, k ∈ {1, 2, 3, 4, 5}. The factors affecting rock blastability are summarized into four indicators as the clustering indicators j, j ∈ {1#, 2#, 3#, 4#}, and the rock mass to be evaluated is regarded as the clustering element i, i ∈ {Ⅰ, Ⅱ, Ⅲ, Ⅳ}.

[0164] Selecting a suitable blastability grade classification method for the lithium ore rock blasting project in the Dangba mining area is very important and directly affects whether the blastability can be accurately applied to guide construction.

[0165] (3) Blasting funnel test.

[0166] 1) Single-hole blasting funnel test.

[0167] The blast holes are arranged on the waist line of the test roadway, that is, 1.2 m above the roadway floor. The rock drill drills 18 (two groups) Φ40 mm blast holes perpendicular to the roadway side, with a hole spacing of 1.5 m. The designed blast hole depths are: 0.40 m, 0.65 m, 0.50 m, 0.55 m, 0.60 m, 0.65 m, 0.70 m, 0.80 m, 0.90 m; all blast holes are arranged perpendicular to the free face.

[0168] To ensure the reliability of the data obtained from the blasting funnel test, the rock powder remaining in the blast holes after rock drilling should be blown clean with water before charging, and the blast hole stemming is strictly carried out according to the design requirements. Therefore, after charging, it is necessary to fill the blast hole with stemming of a length that meets the safety requirements and tamp it firmly. Commonly used is a 1:3 mixture of sand and clay stemming, with a humidity of 18% - 20%. This kind of stemming has both good plasticity and a large friction coefficient. The blasting uses digital electronic detonators for initiation, and each time one hole is blasted. Through the single-hole blasting funnel test, the optimal blasting funnel depth is sought;

[0169] 2) The purpose of conducting the variable hole-spacing multi-hole simultaneous blasting funnel test is to study the blasting situation of the ore and rock at the bottom of the blasting funnel formed when the hole spacing between two adjacent blast holes changes, so as to determine the reasonable range of the hole bottom distance and provide a basis for the design of the blast hole pattern parameters in the subsequent deep stope blasting. The optimal blasting funnel depth obtained from the single-hole blasting funnel test is selected as the charging depth, and the variable hole-spacing simultaneous blasting funnel test is carried out to calculate the unit explosive consumption q under the condition of simultaneous blasting. During the variable hole-spacing simultaneous blasting test, the designed hole spacings are multiples of the optimal blasting funnel radius, and multi-hole simultaneous blasting is carried out with the blast holes perpendicular to the roadway side.

[0170] 3) According to the blasting principle, when the minimum resistance line is less than or equal to the radius of its damage zone during the blasting of a charge, the rock will be thrown or blasted, while when the resistance line is greater than its damage radius. The charge only produces internal blasting and cannot blast the rock from the charge to the free face. Therefore, the resistance line is an important parameter for blasting. Based on this, the parameters of the drawdown horizontal medium-deep hole blasting, ore gathering trough blasting, and eastern boundary deep hole blasting are designed.

[0171] (4) Cut blasting parameters.

[0172] a. Determination of the range of the explosion stress wave fracture zone. Cut blasting is carried out under the condition of only one free face, and the rock breaking is more difficult. Its rock breaking effect can be regarded as the superposition of the blasting effects of each cut hole. The charges in the cut holes detonated simultaneously form radial and circumferential fractures around the blast holes and intersect in the cut cavity to form a spatial fracture network, cutting the rock into fragments and throwing them out under the expansion action of the detonation gas. Therefore, to ensure sufficient rock fragmentation, the layout circle diameter of the cut holes must meet the design requirements.

[0173] b. Deep-hole linear cut blasting utilizes the space of the empty holes to provide a free face and additional space for the blasted rock. Therefore, the reserved space for each cut blasting should meet the swelling requirements of the rock. Determine the blasting additional space requirements of the ore and rock through numerical simulation and on-site tests to ensure the fragmentation effect of the ore and rock. Determine the number and layout pattern of the intermediate empty holes through on-site tests.

[0174] (5) Numerical simulation technology.

[0175] Based on the above research results, determine the blasting schemes for different blasting areas, conduct industrial experiments, count the blasting effects, and verify and obtain the optimal blasting parameter sets.

[0176] 1) Research on the optimization of roadway tunneling blasting parameters and technology.

[0177] a. First of all, it is necessary to require the on-site construction personnel to construct according to the accurate hole layout positions to ensure that the qualification rates of the cut holes and the perimeter holes constructed meet the design requirements.

[0178] b. Optimize the blast hole initiation sequence. Through the analysis of the current initiation network, there are unreasonable parts in the blast hole initiation sequence, which are prone to problems such as blasting clamping. It is necessary to adjust the blast hole initiation sequence and delay time, etc. For the blast holes, initiate in the initiation sequence in the way of forming a small free face to transfer the large free face.

[0179] c. According to the basic theory, the blastability classification of the ore and rock, and the research results of numerical simulation, conduct research on the optimization of the parameters such as the hole depth, angle, and hole bottom distance of the tunneling blast holes, reduce the probability of adverse blasting effects, and ensure production safety.

[0180] The laboratory of a certain university is equipped with relatively complete test equipment, including DH3817 dynamic and static strain test systems, RSM-SY5(T) non-metallic blasting vibration wave detectors, digital ultrasonic blasting vibration wave flaw detectors (CTS-2000Plus), IDT high-speed cameras, presses, three-dimensional dynamic and static combined SHPB and other experimental and test equipment, which can provide support for the static load and cyclic dynamic load combined impact mechanics test of the present invention. Based on the measured physical and mechanical parameters and blastability levels of different ores and rocks, different blasting parameter combinations are studied for the various blasting operation requirements in different regions and environments, which can provide systematic support for mine blasting operations;

[0181] By studying the influence of relevant parameters such as the specific charge of explosives, the minimum resistance line, and the decoupling coefficient on the blasting effect, and establishing the corresponding calculation models; while enriching the theory of the movement of fragmented rocks, it can guide the development of parameter optimization work.

[0182] The blasting parameter design system of the present invention can reduce the economic input of enterprises in underground mine blasting, lower the production cost, improve the operation efficiency, enhance the safety and economy. It has important theoretical and practical significance for ensuring the safe production of mines and improving the economic benefits.

[0183] (6) Determination of material constitutive model and parameters

[0184] a. Rock mass constitutive and parameters: The rock material model adopts constitutive models such as "*MAT_PLASTIC_KINEMATIC" or "MAT_HJC", and the specific parameters will be determined according to the static and dynamic mechanical test results of the rock after the project is carried out.

[0185] b. Explosive material model: In ANSYS / LS-DYNA, the explosive material model adopts the high-energy explosive material model "*MAT_HIGH_EXPLOSIVE_BURN", and the JWL equation of state is used to simulate the explosion of the explosive during the blasting process. The parameters are shown in Table 1.

[0186] Table 1 Parameters of explosive material model

[0187]

[0188] The JWL equation of state relationship during explosive explosion is:

[0189]

[0190] In the formula, P is the pressure, ω, R1, R2 are all material constants, V is the relative volume, and E0 is the internal energy per unit volume; the parameters are shown in Table 2.

[0191] Table 2 Parameters of JWL equation of state

[0192] Mat A / (Gpa) B / (Gpa) <![CDATA[R1]]> <![CDATA[R2]]> ω E / (Gpa) <![CDATA[V0]]> Explosive 214 1.82 4.16 0.96 0.3 4.192 0

[0193] c. Air material model

[0194] The air material model adopts the blank material "*MAT_NULL", and its parameters are shown in Table 3.

[0195] Table 3 Parameters of air material model

[0196] Mat <![CDATA[ρ (kg·m -3 )]]> pc mu terod cerod ym pr Rock 1.29 0 0 0 0 0 0

[0197] The LINEAR-POLYNOMIAL equation of state of air fluid is:

[0198] P = C0 + C1μ + C2μ 2 + C3μ 3 +(C4 + C5μ + C6μ 2 )E0

[0199]

[0200] Wherein, C0, C1, C2, C3, C4, C5, and C6 are all real constants, ρ is the standard density, ρ0 is the reference density, and E0 is the internal energy per unit volume; the parameters in the formula are shown in Table 4.

[0201] Table 4 Parameters of the LINEAR - POLYNOMIAL equation of state

[0202] Mat <![CDATA[C0]]> <![CDATA[C1]]> <![CDATA[C2]]> <![CDATA[C3]]> <![CDATA[C4]]> <![CDATA[C5]]> <![CDATA[C6]]> <![CDATA[E0]]> <![CDATA[V0]]> Air <![CDATA[-1E -6 > 0 0 0 0.4 0.4 0 <![CDATA[2.5E -6 > 0

[0203] Example 2, as Figure 2 shown, the blasting parameter screening system for one - time shaft - forming blasting of blind raises provided by the embodiments of the present invention includes:

[0204] The cut - pattern determination module 1 is used to select a cut - pattern suitable for lithium ore rock according to the mining operation equipment and the results of numerical simulation calculation and analysis, conduct on - site tests, and determine the parameters of the cut - pattern form, the number of empty holes, the hole - distance between the empty holes and the charged holes, and the stemming length;

[0205] The decoupling - coefficient determination module 2 is used to determine the charge amount and the decoupling coefficient of the cut holes and the perimeter holes through numerical simulation calculation and the results of blasting - crater tests;

[0206] The hole - interval delay determination module 3 is used to calculate the influence time of the hole - interval delay on the superposition effect of the explosion stress waves by means of numerical simulation, so as to realize the damage of the stress waves to the rock;

[0207] The stemming and charging module 4 is used to determine the charging method to ensure the accuracy of the charge amount and the position of the explosive, and set a reasonable stemming to ensure that the explosive energy fully acts on the rock mass to be blasted.

[0208] As mentioned above, only the relatively optimal specific implementation manners of the present invention are described, but the protection scope of the present invention is not limited thereto. Any person skilled in the art within the technical scope disclosed by the present invention, any modifications, equivalent substitutions, and improvements made within the spirit and principle of the present invention should all be covered within the protection scope of the present invention.

Claims

1. A method for screening blasting parameters in the one-time shaft sinking blasting of a blind raise, characterized in that The method includes the steps of: S1. Determine the cut - hole pattern. According to the mining operation equipment and the results of numerical simulation calculation and analysis, select a cut - hole plan suitable for lithium ore rock and conduct on - site tests to determine the parameters of the cut - hole form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length. S2. Calculate the charge amount. Determine the charge amount and decoupling coefficient of the cut - hole and the perimeter holes through numerical simulation calculation and the results of blasting funnel tests. S3. Determine the delay time. Use numerical simulation to calculate the influence time of the hole - to - hole delay on the superposition effect of the explosion stress wave. S4. Stemming and charging. Determine the charging method and set reasonable stemming to ensure that the explosive energy fully acts on the rock mass to be blasted. In step S2, the method for determining the decoupling coefficient includes: According to the ratio of the initially determined hole diameter to the cartridge diameter, preset the corresponding decoupling coefficient. After the explosive material model of decoupled charging is determined, use the established JWL equation of state for explosion action and the LINEAR - POLYNOMIAL equation of state of air fluid to calculate whether the tangential tensile stress generated by the excitation of the explosive material model of decoupled charging is greater than the tensile strength of the rock. Optimize the parameters of the directional crack generated by the excitation of the explosive material model of decoupled charging with the decoupling coefficient that meets the energy resolution. Calculating whether the tangential tensile stress generated by the excitation of the explosive material model of decoupled charging is greater than the tensile strength of the rock includes: (1) During the explosion process, simplify the explosive material model into an incompressible Newtonian explosive fluid. According to Newton's second law, obtain the motion differential equation of the explosive fluid. According to the continuity equation derived from the mass conservation equation in the volume element, obtain the formula for the instantaneous tangential tensile stress of the explosive fluid, and calculate the distribution of the tangential tensile stress of the fluid in the cut - hole should be coupled with the flow field and the pressure field. (2) After the explosion of the explosive material model, the tangential tensile stress generated during the conversion to air fluid is determined by the tangential tensile stress conduction equation. (3) During the generation process of explosion cracks, simplify the rock mass into a rigid deformation model. Before the rock mass reaches the fracture strength, the tangential tensile stress and strain generated by the excitation of the explosive material model show a linear relationship. After the stress reaches the fracture strength of the rock mass, it remains constant, and the final strain is the sum of the fracture strain and the crack strain.

2. The method for screening blasting parameters in the one-time shaft sinking blasting of a blind shaft according to claim 1, characterized in that, In step S2, the decoupling coefficient is the ratio of the hole diameter to the cartridge diameter; the value of the decoupling coefficient is 1.5 - 3.0; the perimeter hole spacing is 10 - 20 times the hole diameter; the hole density coefficient of the perimeter holes is taken as 0.8 - 1.0; the charge concentration of the perimeter holes is 70 - 120 kg / m in soft rock, 100 - 150 kg / m in medium - hard rock, and 150 - 250 kg / m in hard rock.

3. The method for screening blasting parameters in the one-time shaft sinking blasting of a blind shaft according to claim 1, characterized in that In step (1), according to Newton's second law, the motion differential equation of the explosive fluid is obtained, and the expression is: where ρ ' is the density of the explosive fluid, o, c, f are the pressure movement velocity components of the explosive fluid along the x, y, and z axes respectively, e is the explosion movement time of the explosive fluid, h is the pressure per unit volume of the explosive fluid, kx, ky, kz are the explosion pressure acceleration components in the three coordinate axis directions respectively, δ is the density of the explosive fluid after the explosion movement, is the Laplace operator.

4. The method for screening blasting parameters in the one-time shaft sinking blasting of a blind shaft according to claim 3, characterized in that, In step (1), according to the continuity equation derived from the mass conservation equation in the volume element, the formula for the instantaneous tangential tensile stress of the explosive fluid is obtained, and the expression is: In the formula, I is the dispersion degree of the explosive fluid. Calculating the distribution of the tangential tensile stress of the fluid in the cut - hole should be coupled with the flow field and the pressure field, and the energy equation is as follows: In the formula, is the slope of the tangential tensile stress - pressure curve of the explosive material model, A is the tangential tensile stress, and J is the pressure conductivity of the explosive material model.

5. The method for screening blasting parameters in the one-time shaft sinking blasting of a blind shaft according to claim 4, characterized in that, In step (2), after the explosive material model explodes and is converted into an air fluid, the tangential tensile stress generated during the excitation process is determined by the tangential tensile stress conduction equation, and the expression is: In the formula, r is the internal energy of the explosive material model.

6. The method for screening blasting parameters in the one-time shaft-sinking blasting of a blind shaft according to claim 5, characterized in that In step (3), the final strain is the sum of the fracture strain and the crack strain, and the expression is: In the formula, is the theoretical stress of the rock mass, E is the fracture modulus, is the theoretical strain of the rock mass, and η s is the stress when fractures occur in the rock mass, is the strength when the rock mass fractures, are the strain corresponding to the fracture of the rock mass and the strain of subsequent fracture generation, respectively.

7. The method for screening blasting parameters in the one-time shaft sinking blasting of a blind shaft according to claim 1, characterized in that, In step S3, determining the delay time includes the steps: S301. Determine the axial height difference D and the straight-line distance l between the pre-blast area of the ore-rock and the blast center of the blast holes in the blast area through the three-dimensional ore-rock structure map of the lithium ore; measure the shear wave velocity u of the blasting vibration wave in the rock mass h and the longitudinal wave velocity u z ; determine the structural influence coefficient G2 and the axial height influence coefficient χ of the blast holes through blasting vibration monitoring; In the same precise delay controlled blasting, the damping ratio of the vibration is the same, and the blasting vibration velocity is affected by the axial height condition of the blast hole. The blasting vibration velocity is expressed as: U (s) = sinθ × G2(D / l) χ Where U (s) is the blasting vibration velocity, and θ is the angle formed by the axial height difference D of the blast hole and the straight-line distance l; S302. According to the blasting vibration velocity calculation formula, the centroid vibration frequency of the lithium ore rock mass conforms to the dimensional analysis theorem and the π theorem. According to the straight-line distance between the pre-blast area of the ore rock and the blast center of the blast hole and the axial height difference of the blast hole, calculate the centroid vibration frequency q of the lithium ore rock mass as: q = F·G2(D / l) χ In the formula, F is the frequency coefficient; S303. By setting the delay time between holes, make the adjacent vibration waveforms differ by N / 4 cycles when reaching the specified blast hole node, where N is an integer, so that the wave peaks of the two waveforms meet when reaching this position, so that the amplitudes are superimposed. The delay time between holes is calculated according to the following formula: According to the calculation formula of the centroid vibration frequency of lithium ore rock mass, let the distances between the pre-blast area of the ore rock and the blast holes be l1, l2…l n , l1 < l2 < … < l n , the axial height differences of the blast holes be D1, D2…D n , the initiation times of the blast holes be s1, s2…s n , s1 < s2 < … < s n ; at the same time, the difference in the distances between the blast holes and the pre-blast area of the ore rock is Δl = l n -l n-1 , the interval time between holes is Δs = s n -s n-1 ; the expression is: Where, l n is the distance between the pre-blast area of ore and rock and the nth blast hole, and l n-1 is the distance between the pre-blast area of ore and rock and the (n - 1)th blast hole, s n is the detonation time of the nth blast hole, and s n-1 is the detonation time of the (n - 1)th blast hole.

8. A blasting parameter screening system for one-time shaft sinking blasting in a blind raise, characterized in that, This system implements the blasting parameter screening method in the one-time shaft forming blasting of a blind raise as described in any one of claims 1-7. This system includes: A cut method determination module (1) for selecting a cut scheme suitable for lithium ore rock according to the mine operation equipment and the results of numerical simulation calculation and analysis, and conducting on-site tests to determine the parameters of the cut form, the number of empty holes, the hole distance between the empty holes and the charged holes, and the stemming length; An uncoupling coefficient determination module (2) for determining the charge amount and uncoupling coefficient of the cut holes and the perimeter holes through numerical simulation calculation and the results of the blasting funnel test; An inter-hole delay determination module (3) for calculating the influence time of the inter-hole delay on the superposition effect of the explosive stress wave by means of numerical simulation to achieve the damage of the rock by the stress wave; A stemming and charging module (4) for determining the charging method to ensure the accuracy of the charge amount and the explosive position, and setting reasonable stemming to ensure that the explosive energy fully acts on the rock mass to be blasted.

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