Calculation of the cross-section of production in the light. Typical cross-sections and determination of the dimensions of the final section of mining exploration workings. (Horizontal workings)

1) The 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 development in the light

Sc \u003d Wsw (+ 0.26 Wsv) \u003d 4900 (1650 +0.29 4900) \u003d 14300 mm2 \u003d 14.3 m2

6) Sichnoy working out in black:

Svch \u003d Vvch (+ 0.26 Vvch) \u003d 5.22 (1.65 + 0.26 5.22) \u003d 15.70 m2

7) Sichnoy of working out in sinking 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. Height of working 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. Wall thickness 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. Area cross section light output, Sc 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 tunnel, Spr 2.1 m2

For horizontal exploration workings, two cross-sectional shapes have been established: trapezoidal (T) and rectangular with a box vault (PS). In fig. 9-10 show typical sections of mine workings of various shapes.

Distinguish between the cross-sectional areas of horizontal workings in the clear, in the sinking and in the rough. Light square (S CB) - this 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 (if any).

The area in the sinking (5 pr) - the area of \u200b\u200bthe mine, which it is obtained in the process of carrying out before the erection of the support, the laying of the rail track, the device of the ballast layer and the laying of engineering communications (cables, air, water pipelines, etc.). Rough area (S BH) - working area, which is obtained by calculation (projected area).

The permissible excess of the area in the penetration over the design (in the rough) are given in table. 2.

table 2

Figure: 9.1. Typical section of trapezoidal workings with wooden lining: a - scraper delivery of rock; b - conveyor delivery of the breed; - manual haulage of the rock; d - locomotive haulage of the rock; d - two-track development with locomotive haulage

Figure: 10. Typical section of workings with monolithic concrete lining with locomotive haulage of rock: a - single-track; b - two-way

Figure: 9.2. Typical cross-section of rectangular-vaulted workings without fastening or with anchor (spray-concrete) fastening: a - scraper delivery of rock; b - conveyor delivery of the breed; - manual haulage of the rock; r - locomotive haulage of rocks; d - two-track development with a locomotive

haulage

Thus, the cross-sectional area of \u200b\u200bthe mine working

or, on the other hand,

Because S B4 \u003d S CB + S Kр, then the calculation of the cross-sectional area of \u200b\u200bthe working begins with the calculation in the light, where S Kp - the section of the working, occupied by the support; K n - the coefficient of the section busting (the coefficient of the excess section - KIS).

The dimensions of the cross-sectional area of \u200b\u200bhorizontal workings in the clear are determined based on the conditions for placing 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 passed with a fastener, and the loader works in a fixed opening. In this case, the calculation is carried out according to the largest dimensions of the rolling stock or loading machine.
2. Development 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):

The decryption of the data is given below (Table 5).

3. Development is carried out without fastening. Then the dimensions of the section are calculated according to the largest dimensions of the tunneling equipment or rolling stock.

Basic dimensions of underground vehicle are standardized for the purpose 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, spray-concrete and combined support.

The main dimensions of the typical sections of mine workings of the T and PS type are given in table. 3 and 4.

Table 3

The main dimensions of the sections of trapezoidal workings (T)

Designation	Section dimensions, mm		Designation	Section dimensions, mm
Designation		Cross-sectional area in the clear, m 2	Designation		Cross-sectional area in the clear, m 2

Table 4

The main dimensions of the sections of the workings are rectangular-vaulted

forms (PS)

Designation	Section dimensions, mm	Cross-sectional area in the clear, m 2
Designation		Cross-sectional area in the clear, m 2

Figure: 11. Schemes of operating conditions of the loading machine in the face: a - in the unsecured bottomhole space; b - in the fixed bottomhole space

Calculation formulas for determining the dimensions of sections of workings of types T and PS are given in table. 5, 6.

Table 5

Trapezoidal formations

	Designation	Calculation formulas
Transport equipment		Selected from catalogs
Free passage
From soil to rail head		h \u003d hi + h p + 1/3 / g shp
Ballast layer (ladder)
Workouts from the rail head		Are selected
to the top		in accordance with PB
Production in the light:
no track
for scraper cleaning
during conveyor delivery of rock		h 4 \u003d h + hi
if there is a rail track:
without ballast		h 4 \u003d h + hi
with ballast		h 4 \u003d h + L3-L2
Rough workings:
without ballast		hs \u003d h 4 + d + ti
with ballast		hs = h 4 + hi + d + ti
Transport equipment		From equipment catalogs
Free passage at height h		Selected in accordance with PB
Aisle level transport equipment
In light at the level of transport equipment:
when scraper cleaning		B \u003d B + 2m
single-track		B \u003d B + m + n
two-way		B \u003d 2B + c + t-n
Workings in the light along the top: without a rail track		b \u003d b-2 (h-H) ctga
if there is a track		B \u003d b- 2 (hi - H) ctga
On the sole:
no track		bi \u003d b + 2 H ctga
in the presence of a track without a ballast layer		Z\u003e 2 \u003d 6 + 2 (# + / ji) ctga
with ballast		Z\u003e 2 \u003d 6 + 2 (# + / ji) ctga

		Designation	Calculation formulas
	Rough workings:
	top base		Bs \u003d b +2 (d+ t 2) sina
	bottom base with ballast layer	Ba	Ba \u003d Bs +2 hs ctga
	without ballast		Ba \u003d b 2 + 2 (d + t 2) sina
	Between transport equipment		Selected according to PB
	eat and mine wall		(t> 250 mm, from> 200 mm)
	Between rolling stock
	Racks, top of round timber
			Estimated

Distance, mm	From the axis of the track (conveyor) to the axis of production: single-track		k \u003d (u + n2 ) -Ы2
Distance, mm	two-way		k \u003d s2 - (u + s2 )
	Cross section: in the light		R\u003d B + 62 + 2L4 / sin a
			Pi \u003d Bs + Ba + 2 / r5 / sin a
	Cross section: in the light		S CB \u003d /24(61 + b 2 ) l2
			S m = /25(63 + 6 4)/2

Table 6

Workings of a rectangular-vaulted shape

		Designation	Calculation formulas

	with sprayed concrete, rod and combined supports		ho \u003d bl4
	with concrete lining		ho \u003d b / 2
	Production in the light:
	without rail track:
	for scraper cleaning		h 4= h + ho
	at conveyor		h 4 \u003d h + /?2 + ho
	in the presence of a track: without ballast layer		h 4 = h+ /?2 + ho
	with ballast		h 4= h + ho
	Rough development		hs= h+ hi + ho +1
	Rough working walls:
	for scraper cleaning
	with ballast layer (ladder)		he = h+ hi
	Transport equipment		Selected from catalogs
	Production in the light:
	single-track		b \u003d B+ m + n
	two-way		b \u003d 2B + c + m + n
	Rough development		bo \u003d b + 2t
	Axial arch arch:
	at ho = N4		R \u003d 0,% 5b
	at ho= S3		R \u003d 0,6926
	Lateral arch of the arch:
	at ho = Nah		r \u003d 0,1736
	at ho = Yb		r \u003d 0.262b
Perimeter transverse development,
	at ho = NA: without ballast		P \u003d 2he + 1,219
	with ballast at ho \u003d b / 3: without ballast		P \u003d 2h + 1,219 P \u003d 2he + 1,33 b
	with ballast		P \u003d 2h + 1,33 b

		Designation	Calculation formulas
Perimeter transverse development,	Roughly: at ho \u003d Ы4at ho \u003d Ы3		/>1=26+1,190 />! = 2*6+1,33 bo
Cross-sectional area of \u200b\u200bthe mine, m 2
	at ho = Nahat ho = S3		S CB \u003d b (h + 0.15b) S CB \u003d b (h + 0.2b)

	without support or rod support		S B4 \u003d b (h 6 + 0, n5b)
	with sprayed concrete and combined lining with concrete lining of a rectangular part of the roadway		S B4 \u003d bo (h 6 + 0.15b)S B h \u003d S CB + S + S 2 + S3 S \u003d 2A 6 / [
	vaulted part of the mine		S 2 = 0.157 (1 + Ao / 6) (6i 2 -6 2)
	subsurface support	S3	Si \u003d 2/27 / + hg (t) -t)
Dimensions of the subsoil part of the support			Chosen according to rock properties and width
			Chosen according to rock properties and width
	Cutting height		working out

All horizontal workings, along which the transportation of goods is carried out, must have at straight sections gaps between the support or equipment placed in the workings, pipelines and the most protruding edge of the rolling stock gauge of at least 0.7 m (n\u003e 0.7) (free passage for people), and on the other hand - not less than 0.25 m (t\u003e 0.25) with wooden, metal and frame structures of reinforced concrete and concrete lining and 0.2 m - with monolithic concrete, stone and reinforced concrete lining.

The free passage width must be maintained at a working height of at least 1.8 m (h = 1,8).

In workings with conveyor delivery, the free passage width must be at least 0.7 m; on the other hand - 0.4 m.

The distance from the upper plane of the conveyor belt to the top or mine roof is at least 0.5 m, and for tensioning and drive heads - at least 0.6 m.

Gap from between opposite electric locomotives (trolleys) along the most protruding edge - not less than 0.2 m (from \u003e 0.2 m).

In the places of trolley coupling and uncoupling, the distance from the support or equipment and pipelines located in the workings to the most protruding edge of the rolling stock gauge must be at least 0.7 m on both sides of the working.

When rolling away by electric contact locomotives, the height of the overhead wire suspension must be at least 1.8 m from the rail head. On landing and loading and unloading sites, at the intersection of workings with workings, where there is a contact wire and along which people move, - at least 2 m.

In the near-shaft yard - in places where people move to the landing site - the suspension height is not less than 2.2 m, in other near-shaft workings - at least 2 m from the rail head.

In the near-shaft yards, on the main haulage workings, in inclined shafts and slopes, when using trolleys with a capacity of up to 2.2 m 3, rails of the R-24 type should be used.

Mine tracks during locomotive haulage, with the exception of workings with heaving soil and with a service life of less than 2 years, must be laid on crushed stone or gravel ballast made of hard rocks with a layer thickness under the sleepers of at least 90 mm.

For horizontal exploration workings, two forms of cross-sections have been established: trapezoidal (T), rectangular-vaulted with a box vault (PS).

Distinguish between 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 lining of the working and its soil, minus the sectional area, which is occupied by the ballast layer poured on the soil of the working.

The area in the sinking (5 P |)) - the area of \u200b\u200bproduction, 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 pipes, etc.). Rough area (5 8H) - the area of \u200b\u200bproduction, 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 working begins with the calculation in the light, where 5 cr is the section of the working occupied by the support; Кп „- the coefficient of the section busting (the coefficient of the excess section - CIS).

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. Development 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 measure it! cross sections calculated
are carried by the largest dimensions of the 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 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 sections of rectangular-vaulted mine 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 is understood here as 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 are manifested at the boundary between a mine working and a rock mass, i.e. within 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 the workings is associated with the redistribution of stresses during their conduction. It manifests itself as;

1) elastic or viscoelastic displacement of rocks without their destruction;

2) collapse (local or regular) in weak, fractured and

finely layered rocks;

3) destruction and displacement of rocks (in particular, collapse) under the influence of limiting stresses in the rock mass along the entire perimeter of the working 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 support, 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 overlying 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 operation.

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, stiffness and stability under static (rock pressure, groundwater pressure, temperature changes, 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.

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 completing 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 sinking sections, hole layout, charge design, work organization schedule).

The sequence of presentation of the material in the explanatory note must comply with the methodological guidelines.

1. CONDITIONS OF EXTRACTION

The working conditions are understood as hydrogeological data and mining 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 CRANK.

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 / m 2 (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 Is 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 M-400 cement 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	Cartridges 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 using electric detonators of instant, 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 gap between the wall of the hole and the cartridges of the explosive, which allows the explosive cartridges to be sent 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 boreholes leads to a deterioration in the delineation of the mine working, excessive destruction of the rock beyond 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 cartridges of small diameter (20 - 22 mm), the expediency of using boreholes of 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 bottomhole, the properties of the blasted rocks, the operability of the explosives used, the type of drilling equipment, the required movement of the bottomhole 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 a hole depth, 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 hole depth:

where: ν - set rate of penetration, m / month;

t c - cycle duration, h;

n with - the number of working days in a month;

n h - the number of working hours per day;

η is the borehole utilization factor (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 commonly 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 ICF, it is necessary to divide the average bottomhole movement by the average depth of the hole.

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 KIP 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 borehole stemming, the order 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 be at least the following values.

Table 4.4.

Table 4.5

Numerical values \u200b\u200bof 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 hole, units;

L zk - length of stemming of contour boreholes (determined according to the table), m;

L to- length of contour bore holes, m.

Table 4.6

Numerical coefficient value K 0

Table 4.7

Reduced stemming length of loop charges L zk / S exp

Coefficient	Linear density of contour borehole loading 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 boreholes, 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 explosive consumption for delineating the development Q to calculated by the 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 and bore 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 conditions of the stressed 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, you should first determine the characteristic value of the borehole length, 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 the 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- conversion factor, taking 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 excavation, the breaking of the main rock volume is carried out under conditions of partial unloading, which allows 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 allows taking into account the type of explosive used.

Table 4.10

with the subsequent delineation of the mine

Coefficient	Depth of work H, m
fortresses		100 - 200	200 - 400	400 - 600
rocks
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 bumpers with preliminary delineation of a mine

Hardness coefficient of rocks	L zb / 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 the conditions of one free surface, to facilitate the operation of cut charges, it is advisable to use one or several 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 assuming 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.):

When further contouring:

(4.30)

When preliminary contouring:

(4.31)

where: P otb- linear density of loading bore holes, kg / m;

e off, e to- 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 holes and the area of \u200b\u200bthe part of the face that falls on these holes is determined by the formulas:

The number of holes designed directly for the destruction of the rock core is determined by the formula (pcs.):

(4.35)

The approximate size of the grid for drilling 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 blasting parameters is compiled in shape.

Drilling and Blasting Parameters Table

Figure: 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 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 holes, the design of the charge in the hole, the cut diagram, the connection diagram of 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 electric 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 amount of 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 guaranteed 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 section, mm 2.

When conducting blasting operations, wires for industrial blasting operations of the VP brand with copper conductors in polyethylene insulation are used as connecting wires and for laying temporary blasting lines. 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 boreholes up to 3m deep in rock with f  6. Drilling is carried out 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- 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 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 various 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 must 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 - number of drilling machines on the rig, pcs;

k 0 - coefficient of simultaneity in machine operation (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 workings is made;

4. Local ventilation fan is selected.

The location of the local ventilation fan (VMP) installation and the direction of the ventilation duct is shown in the “Ventilation certificate”. The passport also indicates the number of VMPs, their type, diameter of the ventilation pipeline, the direction of fresh and outgoing ventilation jets, and security 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 productivity and pressure and lay air ducts with pipes of larger diameter. 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 pollution 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 a rolling stock with internal combustion engines is operating 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.

Figure: 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 the 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 pipes 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

The cross-sectional shape of a horizontal mine workings depends mainly on the type of rock support used to protect the workings from destruction under the pressure of the surrounding rocks and to maintain the required cross-sectional area for the entire period of exploration work. When carrying out workings, they are given a trapezoidal or rectangular sectional shape. The trapezoidal shape is used with timber lining and the presence of slight pressure from the surrounding rocks. The rectangular-vaulted form is used for monolithic concrete, sprayed concrete, anchor and combined (anchor with sprayed concrete) lining and in workings that do not have a lining (for strong stable rocks).
There are cross-sectional areas in the clear, in the rough and in the penetration. The cross-sectional area in the clear is determined by the dimensions of the excavation to the support, minus the areas occupied by the ballast layer of the rail track and the ladder of the footpath. Rough sectional area is the projected area (in penetration). The actual cross-sectional area of \u200b\u200bthe excavation in the sinking is slightly larger than the cross-sectional area in the rough. When driving, it is necessary to observe that the cross-sectional area of \u200b\u200bthe excavation corresponds to the existing "Norms for the excess of the cross-sections of mining exploration openings in the drive in comparison with the sections in the rough during the performance of geological exploration". Depending on the hardness of the rocks, an increase in the sectional area in the rough by a factor of 1.04-1.12 is allowed. A large value of the coefficient corresponds to a cross-sectional area of \u200b\u200b4 m2 in hard rocks.
The size of the cross-section in the clear depends on the purpose of the mine and is determined by the dimensions of the rolling stock and the number of rail tracks, the width of the conveyor, scraper or loading and transport vehicle, taking into account the necessary clearances between these machines and the support, which are regulated by safety rules. The gap between the rolling stock and the support on long sections of the mine working with rail transport is not less than 200 mm with monolithic concrete, anchor and sprayed concrete support and not less than 250 mm with other types of support - flexible metal and wood. If the rolling of the trolleys along the development is carried out manually, then with all types of support this gap is 200 mm.

It might be helpful to read: