Cross-sectional area in penetration. Substantiation of the method of tunneling workings, equipment, shape and dimensions of the cross-section of workings in the light. Methods and means of conducting tunneling works
FEDERAL FISHERIES AGENCY
FEDERAL STATE EDUCATIONAL INSTITUTION
HIGHER EDUCATION
"MURMANSK STATE TECHNICAL UNIVERSITY”
Apatity branch
Department of Mining
CARRYING OUT MINING
Methodical instructions for the implementation of the course project
for students of the specialty
130400 "Mining"
GENERAL ORGANIZATIONAL AND METHODOLOGICAL INSTRUCTIONS
The course project is the final stage of studying the discipline "Carrying out mine workings" and should contribute to the consolidation of theoretical knowledge in the specialty.
The purpose of the course project is to study the technical, technological and organizational issues of driving the projected mine.
When performing the course work, technical, technological and organizational issues of driving the projected mine must be worked out, and the decisions made must ensure the safety of the work.
When working on a term paper, it is necessary to use educational literature, uniform safety rules for mining operations (EPB), as well as materials from domestic and foreign scientific journals.
The explanatory note of the term paper should contain all the necessary calculations and justifications for the decisions made, sketches and diagrams (ventilation scheme, design and tunneling sections, hole layout, charge design, work organization schedule).
The sequence of presentation of the material in the explanatory note should comply with the methodological guidelines.
1. CONDITIONS OF EXTRACTION
The working conditions are understood as hydrogeological data and mining and technical conditions in which the working will be carried out. This section should describe, if they are not specified, the physical and mechanical properties of rocks in terms of their stability, strength, conditions of occurrence and water inflow into the workings during its implementation.
2. METHODS OF PASSAGE AND MECHANIZATION OF WORKS
The method of tunneling used should be the most rational from the point of view of work safety and mechanization of production processes.
When choosing a method of tunneling and means of mechanization of work, it is preferable to use equipment complexes, which to a greater extent provide mechanization of the processes of the tunneling cycle of works.
3. DETERMINATION OF THE SIZE OF THE CROSS-SECTION OF THE EXTRACT AND CALCULATION OF THE CRAFT.
Support calculation.
The load on the support referred to 1 m 2 of the working, with a uniformly distributed disturbed zone, is determined by the formula:
Kgf / m2 (3.29)
where: ρ – volumetric weight of the rock, kg / m 3;
l n - the size of the disturbed zone, m.
The size of the violated zone is determined by the formulas:
a) for workings outside the zone of influence of cleaning works:
b) for inlet and delivery workings:
where: I T - the intensity of the gently dipping small-block system of cracks, pcs / m. running. (Table 1);
K C - coefficient of the state of production (taken equal to 1).
Table 1
table 2
Table 3
Specific adhesion of a rod to concrete and a concrete column to a rock, kgf / cm 2
| Strength indicators | Material name | Fixing mortar on cement M-400 at the age of 28 days. with the composition of the mixture C: P | Mortar on alumina cement M-400 aged | |||||
| 3 days with the composition of the mixture C: P | 12 h at C: P | |||||||
| 1:1 | 1:2 | 1:3 | 1:1 | 1:2 | 1:3 | 1:1 | ||
| Periodic steel | ||||||||
| Smooth round steel | ||||||||
| Concrete post with apatite ore | ||||||||
| Concrete post with oxidized ore | ||||||||
| Concrete post with empty rocks lying side |
The distance between the rods with a square grid of their location is taken from the conditions for preventing stratification and collapse of rocks under the influence of their own weight within the fixed strata according to the formula:
, m (3.40)
where: K zap - safety factor;
m at- coefficient of operating conditions of the rod support (1 - for rods with pre-tension; 2 - for rods without pre-tension).
Table 4.1
Table 4.2
Explosive characteristic
| BB name | Density of explosives in cartridges, g / cm 3 | Efficiency, cm 3 | Detonation velocity, km / s | Type of packaging |
| BB, used in the faces not hazardous for gas or dust | ||||
| Ammonite 6GV | 1,0–1,2 | 360–380 | 3,6–4,8 | Chucks with a diameter of 32, 60, 90 mm |
| Ammonal-200 | 0,95–1,1 | 400–430 | 4.2–4,6 | Cartridges with a diameter of 32mm |
| Ammonal M-10 | 0,95–1,2 | 4,2–4,6 | Also | |
| Ammonal rocky No. 3 | 1,0–1,1 | 450–470 | 4,2–4,6 | Chucks with a diameter of 45, 60, 90 mm |
| Ammonal rocky No. 1 | 1,43–1,58 | 450–480 | 6,0–6,5 | Chucks with a diameter of 36, 45, 60, 90 mm |
| Detonite M | 0,92–1,2 | 450–500 | 40–60 | Cartridges with a diameter of 28, 32, 36 mm |
| BB, used in gas or dust hazardous faces | ||||
| Ammonite AP-5ZhV | 1,0–1,15 | 320–330 | 3,6–4,6 | Cartridges with a diameter of 36 mm |
| Ammonite T-19 | 1,05–1,2 | 267–280 | 3,6–4,3 | Also |
| Ammonite PZhV-20 | 1,05–1,2 | 265–280 | 3,5–4,0 | Also |
In the practice of tunneling works, the most widespread is electric blasting with the help of electric detonators of instantaneous, short-delayed and delayed action, as well as non-electric blasting systems (Nonel, SINV, etc.).
Table 4.3
K zsh values \u200b\u200bfor horizontal workings
Borehole diameter.The diameter of the holes is determined on the basis of the diameter of the explosive cartridges and the required clearance between the wall of the hole and the cartridges of the explosive, which makes it possible to send the explosive cartridges into the hole without effort. Cutters and bits wear out during drilling and sharpening, as a result of which their diameter decreases. Therefore, the initial diameter of the incisors and crowns is used slightly larger than required, and it is 41 - 43 mm for explosive cartridges with a diameter of 36 - 37 mm and 51 - 53 for explosive cartridges with a diameter of 44 - 45 mm. The borehole diameter should be 5–6 mm when the firing cartridge is located first from the borehole mouth, and 7–8 mm when the firing cartridge is not the first from the borehole mouth.
An increase in the diameter of the holes leads to an increase in the explosive charge placed in them, and, consequently, the number of holes decreases. At the same time, an increase in the diameter of the bore-holes leads to a deterioration in the delineation of the mine working, excessive destruction of the rock outside the design contour, and also negatively affects the rate of drilling - the drilling speed decreases.
With an increase in the diameter of the blast-hole charge on the working contour, the zone of destruction of the massif increases and, therefore, the stability of the rocks decreases. Therefore, with a decrease in the cross-section of the mine, it is more expedient to use small-diameter boreholes. With a decrease in the mine section and an increase in the rock hardness, the diameter of the holes and charges, all other things being equal, should decrease. Since the explosives (detonites) produced at the present time are capable of detonating at a high speed in small-diameter cartridges (20-22 mm), the expediency of using bore-holes of a reduced diameter is obvious. And when using explosives with a low detonation velocity of the ammonite type, it is advisable to place cartridges with a diameter of 32 - 40 mm in the boreholes.
Borehole depth. The depth of boreholes is a parameter of tunneling operations, which determines the volume of basic operations in the tunneling cycle and the speed of excavation.
When choosing the depth of boreholes, the area and shape of the bottom hole, the properties of the blasted rocks, the operability of the explosives used, the type of drilling equipment, the required movement of the bottom hole for the explosion, etc. are taken into account. an integer number of driving cycles.
With a shallow (1 - 1.5 m) hole depth, the time of auxiliary work referred to 1 m of bottomhole movement increases (airing, preparatory and final operations when drilling holes and loading rock, loading and blasting explosives, etc.).
With a large (3.5 - 4.5 m) hole depth, the speed of drilling the holes decreases and, ultimately, the relative duration of 1 m of mining increases.
In addition, when choosing the depth of the hole, it should be borne in mind that when blasting at great depths from the earth's surface, where the blasted rocks are compressed from all sides by rock pressure, the destructive effect of the explosion is significantly reduced.
The depth of the holes is determined based on the specified technical rate of penetration, the number and productivity of the mining equipment or according to the production rates.
Knowing the given ROP, you can calculate the depth of the hole:
where: ν - set rate of penetration, m / month;
t c - cycle duration, h;
n с - the number of working days in a month;
n h - the number of working hours per day;
η is the utilization rate of the borehole (BWR).
Borehole utilization factor. The hole utilization rate is the ratio of the used hole depth to the original depth. When explosive charges explode in boreholes, the rock does not break off to the entire depth of the bore-holes, part of the bore-hole is not used in depth and remains in the hearth mass, which is usually called a glass.
To determine the CIP for the entire set of holes, it is necessary to measure the depth of all holes and determine the average depth of the hole. After the explosion of charges, it is necessary to measure the depth of all glasses and determine the average depth of the glass, according to which the average value of the ICF can be found. Therefore, in order to determine the average value of the KIP, it is necessary to divide the value of the average bottomhole movement by the average depth of the borehole.
where: l s - the length of the charge of the hole;
l w - borehole depth.
If bottomhole advance per cycle is specified, then the average borehole depth can be determined by dividing bottomhole advance per cycle by the average KIR value.
The value of the ICF depends on the strength, fracturing and layering of the blasted rocks, the face area, the number of open surfaces in the blasted rock mass, the explosives operability, the depth of the holes, the quality of blasting holes, the sequence of blasting charges and other factors. With the correct determination of all parameters, strict implementation of the blasting technology, the value of the KIR should not be less than the following values.
Table 4.4.
Table 4.5
Numerical values \u200b\u200bof the exponent γ
| cc, kg / m 3 | ||||||||
| , units | 1.843 | 1.892 | 1.940 | 1.987 | 2.033 | 2.125 | 2.214 | 2.301 |
cv - volumetric weight of explosives in charge, kg / m 3
The distance between the contour charges is determined by the formula (m):
(4.6)
where: K 0 - a numerical coefficient that takes into account the interaction of neighboring contour charges and energy losses for the expansion of detonation products in the volume of the borehole, units;
L zk - length of stemming of contour boreholes (determined according to the table), m;
L to- length of contour boreholes, m.
Table 4.6
The value of the numerical coefficient K 0
Table 4.7
Reduced stemming length of loop charges L zk / S exp
| Coefficient | Linear loading density of contour boreholes P to, kg / m | ||
| rock fortresses | 0.4 | 0.5 | 0.6 |
| 4-6 | 0.110-0.097 | 0.121-0.110 | 0.129-0.119 |
| 7-9 | 0.092-0.082 | 0.106-0.097 | 0.115-0.108 |
| 10-14 | 0.077-0.061 | 0.093-0.079 | 0.105-0.092 |
| 15-18 | 0.057-0.046 | 0.076-0.067 | 0.089-0.081 |
| 19-20 | 0.042-0.039 | 0.064-0.061 | 0.079-0.076 |
The approach ratio of contour holes is determined by the formula:
(4.7)
When centuries \u003d 900 - 1100 kg / m 3 this formula can be used in the following form:
(4.8)
Accordingly, the line of least resistance of contour holes is determined by the formula (m):
The number of contour holes is determined by the formula (pcs.):
(4.10)
where: P - full perimeter of the working face, m;
IN - working width at soil level, m
The area of \u200b\u200bthe part of the face that falls on the contour row is (m 2):
(4.11)
To improve the quality of working out the rock at the level of the end parts of the contour holes, an additional charge with a weight equal to (kg) should be placed in the bottom of the latter:
The amount of explosives per contour deflector is determined by the formula (kg):
When preliminary contouring the specific consumption of explosives is determined taking into account the depth of work H(m) by the formula (kg / m 3):
(4.14)
It should be borne in mind that with a decrease in the depth of work, the value q to should not be less than the value determined by formula (4.3).
The distance between the contour holes is calculated according to the formula (4.6), while the value L zk determined by the table (4.8).
Table 4.8
Reduced length of loop charges with preliminary delineation of production
| Coef-nt | Depth of work H, m | |||||||||||
| fortresses | less than 100 | 100-200 | 200-400 | 400-600 | ||||||||
| rocks, f | Linear density of contour borehole loading Р к, kg / m | |||||||||||
| 0.4 | 0.5 | 0.6 | 0.4 | 0.5 | 0.6 | 0.4 | 0.5 | 0.6 | 0.4 | 0.5 | 0.6 | |
| 4-6 | 0.109 | 0.120 | 0.128 | 0.120 | 0.130 | 0.137 | 0.132 | 0.139 | 0.145 | 0.142 | 0.148 | 0.152 |
| 7-9 | 0.093 | 0.106 | 0.116 | 0.106 | 0.117 | 0.125 | 0.118 | 0.128 | 0.135 | 0.130 | 0.138 | 0.144 |
| 10-14 | 0.074 | 0.091 | 0.103 | 0.089 | 0.103 | 0.113 | 0.104 | 0.115 | 0.124 | 0.118 | 0.127 | 0.135 |
| 15-18 | 0.057 | 0.077 | 0.090 | 0.073 | 0.090 | 0.101 | 0.089 | 0.103 | 0.113 | 0.105 | 0.117 | 0.125 |
| 19-20 | 0.046 | 0.067 | 0.082 | 0.062 | 0.081 | 0.093 | 0.080 | 0.096 | 0.106 | 0.097 | 0.110 | 0.119 |
The weight of the additional charge in the bottom hole of the contour holes is determined by the formula (kg):
Number of contour holes N to and the consumption of explosives for delineating the development Q to calculated by formulas. (4.10) and (4.13)
After determining the parameters of contour blasting, they proceed to the calculation of the parameters of loading and placement of cut-and-hammer holes. The basis of the calculation is the value of the specific consumption of explosives for crushing the rock within the drilled volume.
During the subsequent delineation, the core of the face is broken off under the conditions of the stress state of the surrounding rock mass, which leads to the need to increase the energy consumption for crushing the rock in the drilled mass. In this case, first, you should determine the characteristic value of the length of fender holes, taking into account the degree of such influence (m):
(4.16)
Depending on the actual length of fender holes L otb, which, as a rule, is determined by the organization of work and the capabilities of drilling equipment, the value of the specific consumption of explosives for crushing is calculated by the formulas (kg / m 3):
When L off L :
(4.17)
When L off L :
(4.18)
where: e c- a conversion factor that takes into account the type and density of the explosive used.
Table 4.9
The value of the coefficients e cc
During preliminary delineation of the working, the breaking of the main rock volume is carried out under conditions of partial unloading, which makes it possible with the length of the jack holes L off L reduce the specific consumption of explosives to the value determined by the formula (4.17)
After determining the specific consumption of explosives, the parameters of the placement of holes in a straight cut are calculated. The value of the specific consumption of explosives in the cut is determined taking into account the overall efficiency of breaking the rock in the working face:
(4.19)
where: N bp - number of cut holes, units;
R vr - linear density of their loading, kg / m;
L vr- length of cut holes, m;
L zb - stemming length, m.
Absolute value L zb determined by the tables below, followed by division by e centuries, which makes it possible to take into account the type of explosive used.
Table 4.10
during the subsequent delineation of the mine workings
| Coefficient | Depth of work H, m | |||
| fortresses | 100 - 200 | 200 - 400 | 400 - 600 | |
| breeds | ||||
| 4-6 | 0.145 | 0.151 | 0.156 | 0.162 |
| 7-9 | 0.137 | 0.143 | 0.149 | 0.156 |
| 10-14 | 0.128 | 0.135 | 0.142 | 0.149 |
| 15-18 | 0.119 | 0.127 | 0.135 | 0.143 |
| 19-20 | 0.113 | 0.122 | 0.130 | 0.139 |
Table 4.11
Reduced length of tamping of bumpholes with preliminary delineation of a mine
| Hardness coefficient of rocks | L sb / S exp |
| 4-6 | 0.145-0.139 |
| 7-9 | 0.136-0.131 |
| 10-14 | 0.129-0.121 |
| 15-18 | 0.119-0.113 |
| 19-20 | 0.111-0.110 |
The cutting area is determined by the formula (m 2):
(4.20)
The amount of explosives in the cut is determined by the formula (kg)
(4.21)
Since in straight cuts rock crushing is carried out under conditions of one free surface, to facilitate the operation of cut charges, it is advisable to use one or more compensation wells, the minimum diameter of which is determined by the formula (m):
(4.22)
Where: W min - the distance from the well to the nearest cut hole, working for this well, m;
d shp - the diameter of the cut hole, m.
Knowing the area of \u200b\u200bthe cut and taking the shape of the cross-section in the form of one or another flat geometric figure, it is possible to determine the dimensions of the cut cross-section and the parameters of the placement of cut holes (Fig. 4.3):
Square:
Slit:
(4.27)
(4.28)
Fig 4.3 Examples of drilling holes in straight cuts.
After calculating the parameters of the cut, they proceed to calculating the parameters of the chipping.
The total number of fender holes (including soil holes) is determined by the formulas (pcs.):
During subsequent contouring:
(4.30)
When preliminary contouring:
(4.31)
where: R otb- linear density of loading bore holes, kg / m;
e ot, e k- conversion factors, respectively, for bump and loop charges.
The distance between the soil boreholes is calculated by the formula (m):
(4.32)
The line of least resistance of soil boreholes is determined by the formula (m):
(4.33)
The number of soil boreholes and the area of \u200b\u200bthe part of the face that falls on these boreholes is determined by the formulas:
The number of holes intended directly for the destruction of the rock core is determined by the formula (pcs.):
(4.35)
The approximate size of the drilling grid for fender holes is determined by the formula (m):
(4.36)
When preliminary delineating the development S to = 0.
The amount of explosives for breaking the rock within the core and soil zones is determined by the formula (kg):
Based on the calculations and the layout of the holes, a summary table of the parameters of blasting operations is compiled in shape.
Drilling and Blasting Parameters Table
Fig. 4.4 Drill hole layout.
a - hole pattern; b - charge design; 1 - explosive cartridge;
2 - electric detonator.
After calculating all the parameters of the drilling and blasting complex, a certificate of drilling and blasting operations is drawn up.
The blasthole passport should contain a layout of the holes (in three projections), indicate the number and diameter of holes, their depth and tilt angles, the number of blasting series, the sequence of blasting, the amount of charges in the holes, the total and specific consumption of explosives, the consumption of detonators, stemming of each hole and the total amount of stemming material for all holes, as well as the ventilation time of the hole.
To clarify the textual part of this section, the note should provide the corresponding diagrams (the layout of the bore-holes, the diagram of the design of the charge in the bore-hole, the diagram of the cut, the diagram of connecting the detonators in the explosive network).
Calculation of an electric explosion network.
With the electric explosion of charges, it is possible to use all known circuits for connecting resistances in a circuit. The choice of the EM connection scheme depends on the number of exploded EMs and the uniformity of their characteristics. When using electrical explosive devices, the resistance of the explosive network is determined and the result obtained is compared with the limit value of the resistance of the circuit indicated in the instrument's passport. When using power and lighting lines, the resistance of the explosive circuit is determined, then the value of the current passing through a separate EM is calculated and this value is compared with the guaranteed value of the current for a trouble-free explosion. For the guarantee current it is accepted - for 100 ED equal to 1.0 A, and when detonating ED in large groups (up to 300 pieces) 1.3 A and not less than 2.5 A when detonating with alternating current.
When connected in series, the ends of the wires of neighboring EDs are connected in series, and the extreme wires of the first and last EDs are connected to the main wires going to the current source.
The total resistance of the explosive circuit when the ED is connected in series is determined by the formula:
, Ohm (4.38)
where: R 1 - resistance of the main wire in the section from the explosive device to the terminals of the explosive circuit in the working face, Ohm;
R 2 - resistance of additional mounting leads connecting the terminal wires of the ED with each other and with the main wire, Ohm;
n 1 - the number of series-connected ED, pcs;
R 3 - resistance of one ED with terminal wires, Ohm.
Wire resistance is determined by the formula:
where: ρ - specific resistance of the conductor material, (Ohm * mm 2) / m;
l - conductor length, m;
S - conductor cross-section, mm 2.
When conducting blasting operations, as connecting wires and for laying temporary blasting lines, wires for industrial blasting operations of the VP brand with copper conductors in polyethylene insulation are used. The wire is produced by single-core VP1 and two-core VP2x0.7.
For the laying of permanent explosive lines, cables of the NGShM brand are intended. The conductors are made of copper wire. The insulation of the current-carrying conductors is made of self-extinguishing polyethylene.
In exceptional cases, in agreement with the Gosgortekhnadzor authorities, wire VP2x0.7 can be used as permanent explosive lines
Table. 4.12
Table. 4.13
Table 4.14
Drilling holes
Drilling of boreholes is carried out with hand drills, rock drills, drilling rigs.
Hand drills - used for drilling bore-holes up to 3m deep in rock with f 6. Drilling is performed directly from hands or from light supporting devices (SER-19M, ER14D-2M, ER18D-2M, ERP18D-2M). Electric core drills are used when drilling in rock with f 10 (SEC-1, EBK, EBG, EBGP-1).
where: n- the number of drilling machines;
k n - machine reliability factor (0.9);
k 0 - coefficient of simultaneous operation of machines (0.8 - 0.9).
The number of drilling machines is determined on the basis of 4 - 5 m 2 of the bottomhole area per one drilling machine.
Perforators - used for drilling boreholes in rocks with f 5 (PP36V, PP54V, PP54VB, PP63V, PK-3, PK-9, PK-50).
Drilling performance is determined by the formula (m / h):
(4.45)
where: k d- coefficient depending on the borehole diameter (0.7 - 0.72 at d w \u003d 45 mm; 1 at d w \u003d 32 - 36 mm);
k p- coefficient taking into account the type of perforator (1.1 for PP63V, PP54; 1 for PP36V);
and- coefficient taking into account the change in drilling speed in different rocks (0.02 at f \u003d 5-10; 0.3 at f \u003d 10-16).
Drilling rigs... Drilling of boreholes is carried out by drilling rigs or mounted drilling equipment mounted on loading machines.
The choice of a drilling rig for drilling holes in a horizontal opening is made taking into account the following factors:
The type of drilling machine must correspond to the hardness of the rocks in the drilled face;
The dimensions of the drilling zone should be greater than or equal to the height and width of the hole to be drilled;
The maximum length of the drilled holes according to the technical characteristics of the drilling machine (installation) must be coordinated with the maximum length of the holes (according to the drilling blasthole passport);
The width of the drilling rig should not be larger than the vehicles used.
Drilling performance is determined by the formula (m / h):
(4.46)
where: n - the number of drilling machines on the rig, pcs;
k 0 - the coefficient of simultaneity in the operation of machines (0.9 - 1);
k n - unit reliability factor (0.8 - 0.9);
t - duration of auxiliary work (1 - 1.4 min / m);
v m - ROP (m / min).
Table 4.5
Drilling speed
Duration of drilling holes (h):
where: t p- preparatory and final work (0.5-0.7 hours).
Ventilation design.
The design of ventilation of underground workings is carried out in the following sequence:
1. The method of ventilation is selected;
2. A pipeline is selected and its aerodynamic characteristics are determined;
3. The calculation of the amount of air required to ventilate the mine workings is made;
4. A local ventilation fan is selected.
The place of installation of the local ventilation fan (VMP) and the direction of the ventilation duct are shown in the “Ventilation certificate”. The passport also indicates the number of VMPs, their type, the diameter of the ventilation pipeline, the direction of fresh and outgoing ventilation jets, and safety zones.
Ventilation methods.
Workings are ventilated by injection, suction or combined methods.
With the injection method, fresh air flows through pipes to the bottom, and polluted air is removed along the mine. The main advantage of this method is effective ventilation of the bottomhole space with a significant lag of the ventilation pipes from the bottom of the bottomhole. In this case, the use of flexible pipes is possible. However, due to the fact that gases are removed along the entire section and along the length of the mine, it is gassed, which leads to the need to install fans of higher capacity and pressure and lay air ducts with larger diameter pipes. This method is most widespread.
With the suction method, poisonous gases do not spread through the mine, but are sucked out through the pipeline, and fresh air enters the bottomhole space along the mine. The main advantage of this method is that with a sufficiently small distance of the end of the pipeline from the borehole, not exceeding the suction zone, the borehole is ventilated much faster than with other methods, and there is no gas contamination in the main portion of the borehole. This method can be used to ventilate workings when the main sources of production hazard emissions are concentrated in the bottomhole zone. The thrust-in method cannot be used when driving workings through gas-bearing rocks, when operating a rolling stock with an internal combustion engine in them, or with other sources of hazardous emissions, dispersed along the length of the mine.
The combined method involves the use of two fans, one of which operates for exhaust, and the other, installed near the face, for injection. This ventilation method combines the advantages of the delivery and suction methods. In terms of airing time, this method is the most effective. The disadvantages of this method is the obstruction of the production with ventilation equipment.
Fig. 5.1 Ventilation schemes for blind workings.
a - injection; b - suction.
1 - fan; 2 - pipeline.
Table 5.1
Coefficient value R 100
| Pipe diameter, | Metallic | Type M | Textual |
| m | |||
| 0.3 | 990.0 | 1284.0 | 481.0 |
| 0.4 | 228.0 | 305.0 | 108.0 |
| 0.5 | 72.8 | 100.0 | 33.0 |
| 0.6 | 25.0 | 40.1 | 12.5 |
| 0.7 | 11.6 | 28.2 | 5.0 |
| 0.8 | 5.8 | 9.3 | 2.5 |
| 0.9 | 3.0 | 5.1 | 1.3 |
| 1.0 | 1.6 | 3.0 | 0.8 |
Aerodynamic drag of the pipeline. The pressure generated by the fan during its operation on the ventilation pipeline is spent on overcoming the frictional resistance and local resistances, as well as on the high-speed pressure when air leaves the pipeline or when it enters it, with suction ventilation.
The aerodynamic frictional resistance of the pipeline is determined by the formula:
, N * s 2 / m 8 (5.2)
Local resistances of ventilation ducts are usually created by elbows, tees, branches and other shaped parts of pipes. Local resistance values \u200b\u200bare shown below.
Table 5.2
Resistance (N * s 2 / m 8
1) Width of production in the light according to the passport "Kryvbas project":
Vsv \u003d 750 + 1350 + 450 + 1350 + 1000 \u003d 4900 mm.
2) Working width in black:
Vvch \u003d 4900 + 2 60 + 200 \u003d 5220 mm.
3) Clearing height:
Нсв \u003d 1850 + \u003d 1850 + 1650 \u003d mm.
where: \u003d B / 3 \u003d 1650
4) Height of working out in black:
Нвч \u003d Нсв + \u003d 3500 + 60 \u003d 3560 mm.
5) Sich production in the light
Sc \u003d Wsw (+ 0.26 Wsv) \u003d 4900 (1650 +0.29 4900) \u003d 14300 mm2 \u003d 14.3 m2
6) Sich production in black:
Svch \u003d Vvch (+ 0.26 Vvch) \u003d 5.22 (1.65 + 0.26 5.22) \u003d 15.70 m2
7) Siceness of working out in sinking of sinking:
Spr \u003d Vvch (1.02 h 1.05) \u003d 15.70 1.05 \u003d 16.48 m2
Cross-section of the projected mine
The main standard sizes of production:
- 1. Working height in the clear, Нсв. 2200mm.
- 2. Rough working height, Нвч. 2230mm.
- 3. Width of working in the light, Vsv. 2200mm.
- 4. Rough working width, Vvch, 2260mm.
- 5. The height of the box vault, hw 1450mm.
- 6. Thickness of the roof support, d0 30cm.
- 7. Thickness of the wall of the support, dс 30cm.
- 8. Large radius of curvature of the box vault, ?? 1522mm.
- 9. Small radius of curvature of the box vault, ?? 576mm.
- 10. Cross-sectional area of \u200b\u200bopenings, Sс 4.4 m2
- 11. Cross-sectional area of \u200b\u200bthe working in the rough, Svch 4.5 m2
- 12. Cross-sectional area of \u200b\u200bthe mine working in the heading, Spr 2.1 m2
For horizontal exploration workings, two forms of cross-sections have been established: trapezoidal (T), rectangular-vaulted with a box vault (PS).
Distinguish the cross-sectional areas of horizontal workings in the clear, in the sinking and in the rough. The clear area (5 SV) is the area enclosed between the working support and its soil, minus the cross-sectional area, which is occupied by the ballast layer poured on the working soil.
The area in the sinking (5 P |)) - the area of \u200b\u200bthe development, which it is obtained in the process of carrying out before the erection of the support, the laying of the rail track and the device of the ballast layer, the laying of engineering communications (cables, air, water pipelines, etc.). Rough area (5 8H) - working area, which is obtained in the calculation (projected area).
Since 5 VCh \u003d 5 SV + 5 kr, then the calculation of the sectional area of \u200b\u200bthe excavation begins with the calculation in the light, where 5 kr is the section of the excavation occupied by the support; Кп „- the coefficient of the section busting (the coefficient of the excess section - CIS).
The dimensions of the cross-sectional area of \u200b\u200bhorizontal workings in the clear are determined based on the conditions for the placement of transport equipment and other devices, taking into account the necessary clearances, regulated by the Safety Rules.
In this case, it is necessary to consider the following possible cases of excavation and section calculation:
1. Development is carried out with fastening and the loading machine works in a fixed working. In this case, the calculation is carried out according to the largest dimensions of the rolling stock or loading machine.
2. Working out is carried out with fastening, but the support lags behind the face by more than 3 m. In this case, the loader works in the unsecured part of the working.
When calculating the dimensions of the cross-sectional area for the largest dimensions of the rolling stock, it is necessary to make a verification calculation (Fig. 11):
t + B + n "\u003e 2nd +2*2+ t+ In with.+ p; H p +th 3\u003e Az +<* + and-
The decryption of the data is given below.
3. Development is carried out without fastening. Then size it up! cross sections calculated
are carried by the largest dimensions of tunneling equipment or mobile
composition.
The main dimensions of underground vehicles are standardized with the goal of typing the sections of workings, the structure of the support and tunneling equipment.
For trapezoidal workings, standard sections have been developed with the use of solid lining, staggered lining, with only the roof tightening and with the roof and sides tightening.
Typical cross-sections of rectangular-vaulted workings are provided without support, with anchor, sprayed concrete and combined support
Rock pressure
Creation of safe conditions for the functioning of underground structures is one of the main tasks of ensuring the stability of mine workings. The technogenic impact of mining on the geological environment leads to its new state. (The geological environment here means the real physical (geological) space within the earth's crust, which is characterized by a certain set of geological conditions - a set of certain properties and processes).
Quantitatively and qualitatively new force fields appear around the geological-geological object as a part of the geological environment, which appear at the boundary between a mine working and a rock mass, i.e. within the insignificant limits of the rock mass surrounding the mine.
The forces that arise in the massif surrounding the mine are called rock pressure. Rock pressure around workings is associated with the redistribution of stresses during their conduction. It manifests itself as;
1) elastic or viscoelastic displacement of rocks without destruction;
2) collapse (local or regular) in weak, fractured and
fine-bedded rocks;
3) destruction and displacement of rocks (in particular, collapse) under the influence of ultimate stresses in the rock mass along the entire perimeter of the mine section or in its individual sections;
4) extrusion of rocks into the working due to plastic flow, in particular from the side of the soil (heaving of rocks).
The following types of rock pressure are distinguished:
1. Vertical - acts vertically on the lining, filling mass and is a consequence of the pressure of the mass of the overlying rocks.
1. Lateral - is a part of the vertical pressure and depends on the thickness of the rocks lying over the working or the developed layer, the engineering-geological characteristics of the rocks.
3. Dynamic - occurs at high speeds of application of loads: explosion, rock bump, sudden collapse of roof rocks, etc.
4. Primary - the pressure of rocks at the time of excavation.
5. Steady-state - the pressure of rocks after the excavation after some time and not changing for a long period of its functioning.
6. Unsteady - pressure that changes over time due to mining, rock creep and stress relaxation.
7. Static - the pressure of rocks, in which inertial forces are absent or very small.
The increasing complexity of the conditions in which the (underground construction) of mine workings is carried out (large depths of development, permafrost, high seismicity, neotectonic phenomena, acceleration and increase in the volume of technogenic impact, etc.), and the level of development of science made it possible to create modern, more close to real methods for calculating rock pressure.
A new scientific direction arose - the mechanics of underground structures. This is a spider about the principles and methods of calculating underground structures for strength, rigidity and stability under static (rock pressure, groundwater pressure, temperature change, etc.) and dynamic (blasting, earthquakes) effects. She develops methods for calculating support structures.
The mechanics of underground structures arose as a result of the development of rock mechanics - a science that studies the properties and patterns of change in the stress-strain state of rocks in the vicinity of a mine, as well as patterns of interaction of rocks with the support of mine workings to create expedient methods of rock pressure control. The mechanics of underground structures operates with mechanical models of the interaction of the lining with the rock mass, taking into account the geological state of the rocks surrounding the mine, and the calculation schemes of the lining.
The analysis of mechanical models and design schemes is carried out using the methods of the theory of elasticity, plasticity and creep, the theory of fracture, hydrodynamics, structural mechanics, strength of materials, theoretical mechanics.
For open mining exploration workings, substantiate the method of driving, the equipment used and, in accordance with the angle of repose of rocks, select and justify the shape and size of the cross-section, taking into account the design depth of development.
For underground mining and exploration workings, substantiate the method of driving and the corresponding mining equipment, select and justify the shape and dimensions of the cross-section of the workings in the open.
Depending on the physical and mechanical properties of rocks, as well as on the basis of the dimensions of transport and technological equipment (electric locomotives, trolleys, loading machines), taking into account the dimensions of the gaps provided for by the safety rules (PB) during exploration, the dimensions of the cross-section of mine workings in the light are determined ... The dimensions of the workings in the sinking are determined taking into account the thickness of the lining and ties, as well as the height of the track device (ballast, sleepers, rails).
Mine workings can be carried out with or without fastening. Wood, concrete, reinforced concrete, metal and other materials are used as fasteners. Sectional shape can be: rectangular, trapezoidal, arched, round, elliptical.
Horizontal and inclined exploration workings have, as a rule, a short service life, therefore, the main type of support is wood, the section shape is trapezoidal. When driving without fastening, the shape of the section is rectangular-vaulted.
For a trapezoidal section of a mine working with rail transport ( fig. one) it is recommended to calculate the cross-sectional area of \u200b\u200bthe mine in the following sequence.
The dimensions (width and height) of the used electric locomotive or trolley (for manual hauling) determine the width of a single-track working in the clear at the level of the edge of the rolling stock:
B \u003d m + A + n`
and the width of the double-track working:
B \u003d m + 2A + p + n`
m - the size of the gap at the edge of the rolling stock, mm(taken equal to 200 - 250 mm);
p - the gap between the compositions, mm (200mm);
n`- the size of the passage for people at the edge of the rolling stock, mm:
n` \u003d n + * ctg ;
n- the size of the passage at a height of 1800 mm from the level of the ballast layer, equal to at least 700 mm;
h -the height of the electric locomotive (trolley) from the rail head, mm;
h a- the height of the track superstructure from the ballast layer to the rail head, equal to 160 mm;
83 0 - the angle of inclination of the racks, adopted by GOST 22940-85 for exploration workings.
Working height from the rail head to the top in the case of using contact electric locomotives (up to the settlement of the support):
h 1 \u003d h kn. + 200 + 100,
h kn.- suspension height of the contact wire (at least 1800 mm);
200mm - the gap between the contact wire and the support;
100mm- the amount of possible settlement of the lining under the influence of rock pressure.
With other types of transport, the height h 1determined by the graphical construction, taking into account the gap Cbetween the transport equipment and the ventilation pipeline: when transporting battery electric locomotives 250 mm, with manual haulage - 200 mm.
When transporting a battery electric locomotive:
h 1 \u003d h + d t + 250 + 100,
where h - electric locomotive height, mm;
d t- diameter of the ventilation pipe, mm.
Height h 1in general, should not be less than the height of the loader with the bucket raised (for PPN-1s, this height is 2250 mm) minus the height of the ballast layer, i.e. h 1 2250 mm.
Opening width in the clear on the ballast layer:
l 2 \u003d B + 2 (h + h a) * ctg ;
Opening width across the roof:
l 1 \u003d B - 2 (h 1 - h) * ctg ;
Working height from the ballast layer to the support after settlement:
h 2 \u003d h 1 + h a;
Cross-sectional area of \u200b\u200bopenings after settlement:
S sv \u003d 0.5 (l 1 + l 2) * h 2;
Rough working width along the roof (when fastening in staggered directions with tightening the sides):
l 3 \u003d l 1 + 2d,
where d -support post diameter (not less than 160 mm).
Working width on the soil in the rough when fastening in staggered direction with tightening the sides:
l 4 \u003d B + 
,
where h in= 320mm- height from the working soil to the rail head:
h c \u003d h a + h b,
where h b -ballast height.
Working height from soil to support (before settlement):
h 3 `\u003d h 3 + 100,
where ... h 3- the height of the excavation from the soil to the upper stand (after precipitation).
Rough working height before settlement in the presence of tightening:
h 4 `\u003d h 3` + d + 50,
where d - diameter of the fastening timber, mm;
50mm - tightening thickness.
Working height after settlement:
h 4 \u003d h 4 `- 100
Cross-sectional area of \u200b\u200bthe working in the rough before settlement:
S 4 \u003d 0.5 (l 3 + l 4) * h 4 `
Vertical draft equal to 100 mm, allowed only with wooden lining.
In the workings, the laying of wooden sleepers and the laying of the track from rails are used P24 for trolleys up to 2 m 3... When carrying out exploratory workings, trolleys are used VO-0.8; VG-0.7and VG-1,2 with a capacity of 0.8, respectively; 0.7; 1,2 m... When manually rolling with trolleys VO-0.8and VG-0.7, as well as AK-2u electric locomotives use rails P18... The sleepers are laid in a ballast layer with a thickness of 160 mmby immersing them in 2/3 of its thickness.
With a rectangular-vaulted shape, the height of the working in the clear is made up of the wall height from the level of the ballast layer and from the height of the vault ( fig. 2).
Rough working height H is defined as the clear height plus the thickness of the lining in the vault with monolithic concrete lining or plus 50 mm with sprayed concrete, anchor (rod) and combined support. The height of the wall from the level of the rail head to the heel of the arch h 1 during transportation by battery electric locomotives, it is determined depending on the height of the electric locomotive. The height of the workings during transportation by contact electric locomotives must satisfy the conditions under which the minimum clearances are provided between the electric locomotive (trolley) and the support, as well as between the pantograph and the support.
The height of the vertical wall from the tapa level to the heel of the arch h 2 \u003d 1800mm... The height of the vault h 0 take depending on the coefficient of rock hardness on the scale of M.M. Protodyakonov.
For monolithic concrete lining with a strength coefficient f =3:9, h 0 \u003d B / 3.
For sprayed concrete and roof bolting and in unsupported workings f 12 , h 0 \u003d B / 3, and at f 12, h 0 \u003d B / 4.
The curve of a three-center (box) vault is formed by three arcs: axial - R and two side ones - r... The radii of the vault depending on its height:
| Arch height | h 0 | B / 3 | B / 4 |
| Axial arc radius | R | 0,692 | 0,905 |
| Side arc radius | r | 0,262 | 0,173 |
Working width design B 1 with concrete lining, it consists of the width of the working in the clear and doubled the thickness of the lining, and in the case of sprayed concrete, anchor and combined lining, it consists of the width of the working in the clear plus 100 mm.
Single-track clear width:
B \u003d m + A + n
Open double-track working width:
B \u003d m + 2A + p + n,
where n \u003d700mm; p \u003d200mm.
Height of the vertical wall of the mine working from the rail head:
h 1 \u003d h 2 - h a \u003d 1800 - 160 \u003d 1640mm.
Rough working width with sprayed concrete and roof bolting:
B 1 \u003d B +2 \u003d B + 100,
where = 50mm - the thickness of the lining, taken in the calculation.
Cross-sectional area of \u200b\u200bthe working in the clear at the height of the vault h 0 \u003d B / 3:
S St. \u003d B (h 2 + 0.26B),
at h 0 \u003d B / 4: S sv \u003d B (h 2 + 0.175B),
where h 2 \u003d1800mm -the height of the vertical wall from the level of the ladder (ballast layer).
Height of the wall from the working soil:
h 3 \u003d h 2 + h b \u003d h 1 + h B.
Light output parameter at h 0 \u003d B / 3:
P B \u003d 2h 2 + 2.33B,
at h 0 \u003d B / 4: . P B \u003d 2h 2 + 2219B
The cross-sectional area of \u200b\u200bthe working out in the rough with sprayed concrete, anchor, combined support with h 0 \u003d B / 3:
S h. \u003d B 1 (h 3 + 0.26B 1),
at h 0 \u003d B / 4: S h \u003d B 1 (h 3 + 0.175B 1).
After determining the cross-sectional area, we take GOST 22940-85 the nearest standard section and write down its dimensions for further calculations. According to this standard, only the cross-sectional area of \u200b\u200bthe working in the clear is determined, and the cross-sectional area is roughly set depending on the adopted cross-sectional shape, type and thickness of the support according to the above formulas.
In the table 1 shows the typical cross-sections and basic equipment adopted for calculating the cross-section in the clear, as well as the dimensions of the basic vehicles.
Pits by depth are conventionally divided into shallow (up to 5 m), medium (5 - 10) and deep (up to 40 m). The depth of the pits depends on the stage of exploration and geological conditions. Depending on the physical and mechanical properties of the rocks, the method of penetration and the structure of the support, the pits are round and rectangular. With increasing pit depth, the clear cross-sectional area increases. Pits up to 10 m usually have one compartment, and with a depth of up to 20 m can be with two branches. Typical sections ( GOST 41-02-206-81), it is planned to drill pits with a clear cross-sectional area from 0.8 to 4 m 3 and geometrical dimensions (Table 2).
The dimensions of the cross-section of horizontal mine workings in the light depend on its purpose and are determined based on the dimensions of the rolling stock and the equipment located in the mine, ensuring the passage of the required amount of air, the gaps between the protruding parts of the rolling stock and the support, provided for by the Safety Rules and the method of movement of people.
In our case, we are designing a horizontal vaulted excavation with roof bolting.
Rectangular-vaulted sections are used when driving workings without support or with the construction of lightweight support structures. The height of the vault in sections from 2 to 6.8 m 2 is?. working width.
The clear cross-sectional area is the area along the inner contour of the support installed in the working
Calculation of the section of the mine
Cutting width
b \u003d b c + 2c \u003d 0.95 + 2 0.3 \u003d 1.55m
where b c - scraper width, m;
c - the gap between the scraper and the side of the mine, m.
In a mine of the type under consideration, people are allowed to walk only when the scraper installation is inoperative. Thus, the clearance height is assumed to be minimal, i.e. 1.8 m.
Arch height
Side cutting height (up to the heel of the arch):
1.8 - minimum production height according to PB
According to the calculated cross-sectional area in the light, the nearest larger of the standard cross-sections from table is taken. 2 (Tutorial "Conducting horizontal exploration workings and chambers" Authors V.I.
A typical cross-section of the production of substations is accepted - 2.7
The main dimensions of the cross-section of the working in the clear:
Working width, mm - b \u003d 1550 mm
Working height to the heel of the arch, mm - h b \u003d 1320 mm
Working height, mm - h \u003d 1850 mm
The radius of the axial arc of the arch, mm - R \u003d 1070 mm
The radius of the side arc of the arch, mm - r \u003d 410 mm
Cross-sectional area of \u200b\u200bopenings, m 2 - S sv \u003d 2.7 m 2.
For workings with roof bolting:
where is the height of the working on the side, taking into account the exit of the anchors along the roof into the working by the value d \u003d 0.05 m.
Calculation of the strong dimensions of the support, drawing up the fastening passport
Due to the small section of the mine working, short service life, mining and geological conditions and available materials, we use metal expansion bolting AR-1
All calculations of the strength of anchoring in the borehole of the roof bolting were made according to the formulas from the reference book "Theory and practice of using roof bolting" Author A.P. Shirokov. Moscow "Nedra" 1981
c - angle of friction of rocks, 30 degrees
D - the diameter of the spacer sleeve, 32cm
h - height of the spacer sleeve, 30cm
y s is the ultimate compressive strength of the rock
b - half the angle of a symmetrical wedge, 2 degrees
p 1 - angle of friction of steel on steel, 0.2 degrees
The required length of the anchor L and in the roof and the height of the possible fallout of the working rocks is found from the expressions:
L a \u003d b + L 2 + L 3 \u003d 0.04 + 0.35 + 0.05 \u003d 0.44m;
where L 2 - the value of the depth of the anchors beyond the contour of the possible fallout of rocks (taken equal to 0.35 m); L 3 - the length of the anchor protruding beyond the mine contour, L k \u003d 0.05 m; and n \u003d half-span of working in driving, m; h is the height of the excavation in the sinking, m.
Coefficient characterizing the stability of the sides of the mine;
Coefficient characterizing the slope of the creep prism in the sides of the working (taken according to Table 12.1. Theory and practice of using bolting. Author A.P. Shirokov. Moscow "Nedra" 1981);
c b - angle of internal friction (resistance) of rocks in the sides of the mine; К к - coefficient taking into account the decrease in the strength of rocks in the roof of the mine (taken according to Table 13.1);
f to - the coefficient of rock hardness in the roof of the workings;
K sr - the coefficient of concentration of compressive stresses on the contour of the mine, the value of which is taken from the table. 12.2;
d - the average specific weight of rock strata overlying the mine to the surface, MN / m 3; Н - working depth from the surface, m;
K b - coefficient taking into account the decrease in the strength of rocks in the sides of the working, the value of which is taken according to Table 12.1;
f b - coefficient of rock hardness according to M.M. Protodyakonov in the sides of the mine.
We accept the length of the anchor in the roof L k \u003d 0.5 m.
Due to the fact that w0, anchoring of the sides of the excavation is not performed.
Roof area supported by one anchor
where F to - the area of \u200b\u200bthe roof, supported by one anchor, m 2;
P k - the strength of the anchor in the hole drilled in the roof;
The safety factor, taking into account the uneven distribution of the load on the anchor and the possibility of surcharge from the overlying layers, is taken equal to 4.5;
b - the angle of inclination of the working, degree 0 0
Distance between anchor in a row:
where L n is the step of installing the anchors along the width of the working, m;
L y - the distance between the rows of anchors, m, taken 1.4 m
Number of anchors in a row
where L b \u003d 1.33b \u003d 1.331.55 \u003d 2.06m is the part of the working perimeter, which is to be anchored along the roof, m. Where b is the width of the working in the rough.
Accepts 2 anchors in a row.
Drawing up a fastening passport.
Clear cut width:
B \u003d B + 2m \u003d 950 + 3002 \u003d 1550mm.
Cutting arch height
h about \u003d b / 3 \u003d 1550/3 \u003d 520mm.
Rough cut height
h 2 \u003d h + h o + t \u003d 1320 + 520 + 50 \u003d 1890mm.
Rough cut wall height
h 3 \u003d h + t \u003d 1320 + 50 \u003d 1370mm.
Radius of the axial arc of the cutting arch
R \u003d 0.692b \u003d 0.6921550 × 1070mm.
Radius of the lateral arc of the cutting arch
r \u003d 0.692b \u003d 0.6921550 × 410mm.
Clear cross-sectional area:
S sv \u003d b (h + 0.26b) \u003d 1.55 (1.32 + 0.261.55)? 2.7m 2
Cross-sectional perimeter of the cut in the light:
P \u003d 2h + 1.33b \u003d 21.32 + 1.331.55 \u003d 4.7m.
Cross-sectional area of \u200b\u200bthe cut in the rough:
S high \u003d b (h 3 + 0.26b) \u003d 1.55 (1.37 + 0.261.55) \u003d 2.75m 2.
Rough cut cross-sectional perimeter:
P \u003d 2h + 1.33b \u003d 21.37 + 1.331.55 \u003d 4.8m
Distance between anchors in a row: b 1 \u003d 1200mm.
Distance between rows of anchors: L \u003d 1.4 m
Depth of holes for anchors: l \u003d 500mm.
Diameter of holes for anchors: \u003d 43mm.
The maximum lag of anchor support from the bottom of the face is taken to be 3 m.
Scheme for calculating the dimensions of the cross-section when using scraper equipment in the development of a rectangular-vaulted section shape.