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

There are cross-sectional areas of horizontal workings in the open, in the tunnel and in the rough. Clear area (S CB) - this is the area enclosed between the excavation support and its soil, minus the cross-sectional area that is occupied by the ballast layer poured on the excavation soil (if any).

The area in the excavation (5 pr) is the excavation area, which is obtained during the process before the construction of support, laying of the rail track, installation of the ballast layer and laying of utilities (cables, air, water pipelines, etc.). Rough area (S BH) - excavation area, which is obtained in the calculation (design area).

The permissible excess of the area in the excavation over the design (draft) is given in table. 2.

table 2

Rice. 9.1. Typical cross-section of trapezoidal workings with wooden support: a - scraper delivery of rock; b - conveyor delivery of rock; c - manual removal of rock; g - locomotive haulage of rock; d - two-track working with locomotive haulage of rock

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

Rice. 9.2. A typical cross-section of rectangular-vaulted workings without fastening or with anchor (sprayed concrete) fastening: a - scraper delivery of rock; b - conveyor delivery of rock; c - manual removal of rock; G - locomotive haulage of rock; d - two-track development with locomotive

rock removal

Thus, the cross-sectional area of ​​the excavation

or, on the other hand,

Because S B4 = S CB + S Kр, then the calculation of the cross-sectional area of ​​the excavation begins with calculation in the open, where S Kp- section of the excavation occupied by the support; K p- coefficient of cross-section enumeration (coefficient of excess section - CIS).

The dimensions of the cross-sectional area of ​​horizontal openings 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 cross-section calculation:

• 1. The excavation is traversed with support, and the loading machine operates in the secured excavation. In this case, the calculation is carried out according to largest dimensions rolling stock or loading machine.
• 2. The working is carried out with support, but the support lags behind the face by more than 3 m. B in this case the loading machine operates in an unsecured part of the excavation.

When calculating the dimensions of the cross-sectional area based on the largest dimensions of the rolling stock, it is necessary to make a verification calculation (Fig. 11):

A breakdown of the data is given below (Table 5).

3. The working is carried out without fastening. Then the cross-sectional dimensions are calculated based on the largest dimensions of the tunneling equipment or rolling stock.

Main dimensions of underground Vehicle standardized for the purpose of typifying the sections of workings, the design of support and tunneling equipment.

For trapezoidal-shaped workings, standard sections have been developed using solid support, staggered support, with only the roof tightened, and with the roof and sides tightened.

Typical cross-sections of rectangular-vaulted workings are provided without support, with anchor, shot-concrete and combined support.

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

Table 3

Main cross-sectional dimensions of trapezoidal workings (T)

Designation	Section dimensions, mm				Designation	Section dimensions, mm
				Clear cross-sectional area, m2					Clear cross-sectional area, m2

Table 4

Main cross-sectional dimensions of rectangular-vaulted workings

forms (PS)

Designation	Section dimensions, mm					Clear cross-sectional area, m2

Rice. 11. Schemes of the operating conditions of the loading machine at the face: a - in an unsecured bottom-hole space; b - in a fixed bottom-hole space

Calculation formulas for determining the cross-sectional sizes of workings of types T and PS are given in Table. 5, 6.

Table 5

Trapezoidal workings

		Designation	Calculation formulas
	Transport equipment		Selected from catalogs
	Free passage
	From soil to rail head		h =hi + h p + 1/3 /g sh
	Ballast layer (ladder)
	Workings from the rail head		Selected
	to the top		in accordance with PB
	Productions in the world:
	without track
	during scraper removal of rock
	during conveyor delivery of rock		h 4 = h + hi
	if there is a rail track:
	without ballast layer		h4 = h + hi
	with ballast layer		h4 = h+ L3-L2
	Rough workings:
	without ballast layer		hs = h 4 + d + ti
	with ballast layer		hs = h 4 + hi + d + ti
	Transport equipment		From equipment catalogs
	Free passage at height h		Selected in accordance with the safety regulations
	Passage at the level of transport equipment
	In light at the level of transport equipment:
	during scraper cleaning		b = b + 2m
	single-track		b = B + t + n
	double track		b = 2B + c + t-p
	Workings in the open along the upper gorge: without a rail track		b = b-2(h-H) ctga
	if there is a rail track		B=b- 2(hi - H) ctga
	On the sole:
	without track		bi = b + 2H ctga
	in the presence of a rail track without ballast layer		Z>2 = 6 + 2(#+/ji)ctga
	with ballast layer		Z>2 = 6 + 2(#+/ji)ctga

		Designation	Calculation formulas
	Rough workings:
	upper base		b3 = b+2 (d+ t 2) sina
	lower base with ballast layer	Ba	Ba = b3 +2 hs ctga
	without ballast layer		Ba = b 2 + 2 (d + t 2) sina
	Between transport equipment		Selected according to PB
	eat and the wall of the excavation		(T> 250 mm, With> 200 mm)
	Between rolling stock
	Racks, top made of round timber
			Estimated
Distance, mm	From the axis of the track (conveyor) to the axis of the excavation: single-track		k = (u + s2 )-Y2
	double track		k = S2 -(u+s2 )
	Cross section: clear		R= b+ 62 + 2Л4/sin a
			Pi = bz+ bа + 2/r5/sin a
	Cross section: clear		S CB = /24(61 +b 2 )l2
			S m = /25(63 + 6 4)/2

Table 6

Rectangular-vaulted workings

		Designation	Calculation formulas
	for shot concrete, rod and combined support		ho = bl4
	with concrete support		ho = b/2
	Productions in the world:
	without track:
	during scraper removal of rock		h 4= h + ho
	with a conveyor		h 4 =h + /?2 + ho
	with a rail track: without ballast layer		h 4 = h+ /?2 + ho
	with ballast layer		h 4= h + ho
	Rough workings		hs= h+ hi + ho +1
	Rough working walls:
	during scraper removal of rock
	with ballast layer (ladder)		he = h+ hi
	Transport equipment		Selected from catalogs
	Productions in the world:
	single-track		b=B+ m+n
	double track		b = 2B + c + m + n
	Rough workings		bo = b + 2t
	Axial arc of the arch:
	at ho = N4		R = 0.%5b
	at ho= Y 3		R= 0,6926
	Lateral arch:
	at ho = YA		r = 0,1736
	at ho = ЪЪ		r = 0.262b
Perimeter transverse production,
	at ho = YA: without ballast layer		P = 2he+ 1,219
	with ballast layer at ho = b/3: without ballast layer		P = 2h+ 1,219 P = 2he + 1,33 b
	with ballast layer		P = 2h+ 1,33 b

		Designation	Calculation formulas
Perimeter transverse production,	Roughly: at ho = Н4 at ho = ы 3		/>1=2*6+1,19*0 />! = 2*6+1,33 bo
Cross-sectional area of ​​the mine, m 2
	at ho = YA at ho = Y 3		S CB = b(h + 0.15b) S CB = b(h + 0.2b)
	without support or rod support		SB4= b(h 6 +0,n5b)
	with shot-concrete and combined lining with concrete lining of a rectangular part of the excavation		SB4= bo(h 6 +0.15b)S B h = S CB + S+ S 2 +S3 S= 2A 6 /[
	vaulted part of the excavation		S 2 = 0.157(1 + Ao/6)(6i 2 -6 2)
	subsoil part of the support	S3	Si = 2/27/ + hg(t)-t)
Dimensions of the subsoil part of the support			Selected depending on the properties of the rocks and width
	Cutting height		production

All horizontal excavations along which cargo is transported must have gaps in straight sections between the support or equipment placed in the excavation, pipelines and the most protruding edge of the rolling stock gauge of at least 0.7 m (n> 0.7) (free passage for people), and on the other side - at least 0.25 m (t> 0.25) for wooden, metal and frame structures, reinforced concrete and concrete support and 0.2 m - for monolithic concrete, stone and reinforced concrete support.

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

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

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

Gap With between oncoming electric locomotives (trolleys) along the most protruding edge - at least 0.2 m (With> 0.2 m).

In places where trolleys are coupled and uncoupled, the distance from the support or equipment and pipelines placed in the excavations to the most protruding edge of the rolling stock gauge must be at least 0.7 m on both sides of the excavation.

When rolling by contact electric locomotives, the height of the contact wire suspension must be at least 1.8 m from the rail head. At landing and loading and unloading areas, at the intersections of excavations with excavations, 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 height of the suspension is at least 2.2 m, in other near-shaft workings - at least 2 m from the rail head.

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

Mine rail 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 strong rocks with a layer thickness under the sleepers of at least 90 mm.

For horizontal mining and exploration workings, two cross-sectional shapes have been established: trapezoidal (T), rectangular-vaulted with a box vault (PS).

There are cross-sectional areas of horizontal workings in the open, in the tunnel and in the rough. The clear area (5 SV) is the area enclosed between the excavation support and its soil, minus the cross-sectional area that is occupied by the ballast layer poured onto the excavation soil.

The area in the excavation (5 P|)) is the area of ​​the excavation, which is obtained during the process before the construction of support, laying of the rail track and installation of the ballast layer, laying of utilities (cables, air, water pipes, etc.). Rough area (5 8H) - the excavation area that is obtained during the calculation (design area).

Since 5 HF = 5 SV + 5 cr, then the calculation of the cross-sectional area of ​​the excavation begins with a calculation in the open, where 5 cr is the cross-section of the excavation occupied by the support; Кп„ - coefficient of cross-section search (coefficient of excess cross-section - KIS).

The dimensions of the cross-sectional area of ​​horizontal openings 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 carrying out excavations and calculating the cross-section:

1. The working is carried out with fastening and the loading machine operates in the fixed working. In this case, the calculation is carried out based on the largest dimensions of the rolling stock or loading machine.

2. The excavation is traversed with support, but the support lags behind the face by more than 3 m. In this case, the loading machine operates in the unsupported part of the excavation.

When calculating the dimensions of the cross-sectional area based on the largest dimensions of the rolling stock, it is necessary to make a verification calculation (Fig. 11):

t + B + p">2nd + 2*2+ T+ In the village+ P; N r + th 3 > Az +<* + And-

A breakdown of the data is given below.

3. The working is carried out without fastening. Then size it up! cross sections calculated
are based on the largest dimensions of tunneling equipment or mobile
composition.

The main dimensions of underground vehicles are standardized with the aim of typifying the sections of workings, the design of support and tunneling equipment.

For trapezoidal-shaped workings, standard sections have been developed using solid support, staggered support, with only the roof tightened, and with the roof and sides tightened.

Typical sections of rectangular-vaulted workings are provided without support, with anchor, shotcrete and combined support

Rock pressure

Creating safe operating conditions for underground structures is one of the main tasks of ensuring the sustainability of mine workings. The technogenic impact of mining on the geological environment leads to a new state. (The geological environment here refers to 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 arise around the geological object as part of the geological environment, which manifest themselves at the boundary between the mine workings and the rock mass, i.e. within small limits of the rock mass surrounding the mine.

The forces arising in the massif surrounding the mine opening are called rock pressure. The rock pressure around the workings is associated with the redistribution of stresses during their construction. It manifests itself in the form;

1) elastic or elastic-viscous displacement of rocks without their destruction;

2) landslide formation (local or regular) in weak, fractured and

finely layered rocks;

3) destruction and displacement of rocks (in particular, rock formation) under the influence of extreme stresses in the massif along the entire perimeter of the excavation section or in its individual sections;

4) extrusion of rocks into the excavation due to plastic flow, in particular from the soil (rock heaving).

The following types of rock pressure are distinguished:

1. Vertical - acts vertically on the support, backfill mass and is a consequence of the pressure of the mass of overlying rocks.

1. Lateral - is part of the vertical pressure and depends on the thickness of the rocks lying above the workings or the layer being developed, and the engineering and geological characteristics of the rocks.

3. Dynamic - occurs at high rates of application of loads: explosion, rock burst, sudden collapse of roof rocks, etc.

4. Primary - rock pressure at the time of excavation.

5. Steady - the pressure of rocks after the excavation has passed after some time and does not change for a long period of its operation.

6. Unsteady - pressure that changes over time due to mining, rock creep and stress relaxation.

7. Static - rock pressure in which inertial forces are absent or very small.

The increasing complexity of the conditions in which mining (underground construction) is carried out (great 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 have made it possible to create modern ones that are closer to real ones. methods for calculating rock pressure.

A new scientific direction has emerged - the mechanics of underground structures. This is a book about the principles and methods of calculating underground structures for strength, rigidity and stability under static (rock pressure, groundwater pressure, temperature changes, etc.) and dynamic (blasting, earthquakes) influences. 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 changes in the stress-strain state of rocks in the vicinity of a working, as well as the patterns of interaction of rocks with the support of mine workings to create appropriate methods for controlling rock pressure. The mechanics of underground structures operates with mechanical models of the interaction of the support with the rock mass, taking into account the geological state of the rocks surrounding the excavation, and design diagrams of the support.

Analysis of mechanical models and calculation schemes is carried out using methods of the theory of elasticity, plasticity and creep, fracture theory, hydrodynamics, structural mechanics, strength of materials, theoretical mechanics.

FEDERAL FISHERIES AGENCY

FEDERAL STATE EDUCATIONAL INSTITUTION

HIGHER EDUCATION

“MURMANSK STATE TECHNICAL UNIVERSITY”

Apatitsky branch

Department of Mining Engineering

MINING WORKING

Guidelines for completing the course project

for specialty students

130400 "Mining"

GENERAL ORGANIZATIONAL AND METHODOLOGICAL INSTRUCTIONS

The course project is the final stage of studying the discipline “Carrying out mining operations” and should help consolidate theoretical knowledge in the specialty.

The purpose of the course project is to study the technical, technological and organizational issues of sinking the designed excavation.

When performing course work, the technical, technological and organizational issues of sinking the designed excavation must be worked out, and the decisions made must ensure the safety of the work.

When working on coursework, it is necessary to use educational literature, uniform safety rules for mining operations (USR), as well as materials from domestic and foreign scientific journals.

The explanatory note of the course work must contain all the necessary calculations and justifications for the decisions made, sketches and diagrams (ventilation diagram, design and excavation cross-sections, drill hole layout, charge design, work organization schedule).

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

1. CONDITIONS FOR PRODUCTION

The conditions of excavation are understood as hydrogeological data and mining-technical conditions in which the excavation 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. CONSTRUCTION METHODS AND MECHANIZATION OF WORK

The excavation method used must be the most rational from the point of view of work safety and mechanization of production processes.

When choosing a tunneling method and means of mechanization of work, it is preferable to use equipment complexes that, to a greater extent, ensure mechanization of the processes of the tunneling cycle of work.

3. DETERMINATION OF THE DIMENSIONS OF THE CROSS SECTION OF THE MINING AND CALCULATION OF THE SUPPORT.

Calculation of support.

The load on the support per 1 m2 of excavation, with a uniformly distributed disturbed zone, is determined by the formula:

Kgs/m 2 (3.29)

Where: ρ – volumetric weight of the rock, kg/m 3 ;

l n– dimensions of the disturbed zone, m.

The size of the disturbed zone is determined by the formulas:

a) for workings outside the zone of influence of mining operations:

b) for inlet and delivery workings:

Where: I T– intensity of gently dipping small-block crack system, pcs/m. linear (Table 1);

K C– production condition coefficient (assumed equal to 1).

Table 1

table 2

Table 3

Specific adhesion of a rod to concrete and a concrete column to rock, kgf/cm 2

Strength indicators	Name of material	Fixing solution on cement M-400 at the age of 28 days. with a mixture of C:P	Aged alumina cement mortar M-400
3 days with a mixture of C:P	12 hours at C:P
1:1	1:2	1:3	1:1	1:2	1:3	1:1
	Periodic steel
Smooth round steel
	Concrete column with apatite ore
Concrete pillar with oxidized ore
Concrete pillar with waste rocks lying side

The distance between the rods with a square grid of their location is taken from the conditions for preventing delamination and collapse of rocks under the influence of their own weight within the fixed thickness according to the formula:

, m (3.40)

Where: K zap– safety factor;

m y– coefficient of operating conditions of the rod support (1 – for rods with pre-tensioning; 2 – for rods without pre-tensioning).

Table 4.1

Table 4.2

Characteristics of explosives

Name of explosive	Density of explosives in cartridges, g/cm 3	Performance, cm 3	Detonation speed, km/s	Type of packaging
BB,used in mines that are not hazardous due to gas or dust
Ammonit 6ZhV	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	Same
Ammonal rock No. 3	1,0–1,1	450–470	4,2–4,6	Chucks with a diameter of 45, 60, 90 mm
Ammonal rock No. 1	1,43–1,58	450–480	6,0–6,5	Chucks with a diameter of 36, 45, 60, 90 mm
Detonit M	0,92–1,2	450–500	40–60	Cartridges with a diameter of 28, 32, 36 mm
BB,used in mines hazardous for gas or dust
Ammonite AP-5ZhV	1,0–1,15	320–330	3,6–4,6	Chucks with a diameter of 36 mm
Ammonite T-19	1,05–1,2	267–280	3,6–4,3	Same
Ammonite PZhV-20	1,05–1,2	265–280	3,5–4,0	Same

In the practice of tunneling work, electric blasting using instantaneous, short-delayed and delayed electric detonators, as well as non-electric blasting systems (Nonel, SINV, etc.), has become most widespread.

Table 4.3

Values ​​of K zsh for horizontal workings

Hole diameter. The diameter of the holes is determined based on the diameter of the explosive cartridges and the required gap between the hole wall and the explosive cartridges, 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 cutters 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 diameter of the hole should be 5–6 mm when the firing cartridge is located first from the mouth of the hole, and 7–8 mm when the firing cartridge is not located first from the mouth of the hole.

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 holes leads to a deterioration in the delineation of the mine workings, excessive destruction of the rock beyond the design contour, and also negatively affects the pace of drilling - the drilling speed decreases.

With an increase in the diameter of the hole charge on the excavation contour, the zone of destruction of the massif increases and, consequently, the stability of the rocks decreases. Therefore, with a decrease in the cross-section of the excavation, it is more expedient to use small-diameter holes. With a decrease in the cross-section of the excavation and an increase in the strength of the rocks, the diameter of the holes and charges, other things being equal, should decrease. Since currently produced explosives (detonites) are capable of detonating at high speed in cartridges of small diameter (20 - 22 mm), the advisability of using drill holes of reduced diameter is obvious. And when using explosives with a low detonation speed such as ammonites, it is advisable to place cartridges with a diameter of 32 - 40 mm in the holes.

Hole depth. The depth of the holes is a parameter of tunneling work that determines the volume of main operations in the tunneling cycle and the speed of excavation.

When choosing the hole depth, the area and shape of the face, the properties of the rocks to be blasted, the performance of the explosives used, the type of drilling equipment, the required movement of the face during the explosion, etc. are taken into account. It is desirable that the duration of the tunneling cycle be a shift or an integer number of shifts, and in case of multi-cycle work, a shift ends an integer number of tunneling cycles.

With a small (1 - 1.5 m) hole depth, the time of auxiliary work per 1 m of face advance increases (ventilation, 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 holes decreases and, ultimately, the relative duration of 1 m of mining increases.

In addition, when choosing the depth of a hole, it should be taken into account that when blasting at great depths from the earth's surface, where the rocks being blasted are compressed on 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 speed of penetration, the number and productivity of mining equipment, or according to production standards.

Knowing the specified penetration speed, you can calculate the hole depth:

where: ν – specified penetration speed, m/month;

t c – cycle duration, h;

n с – number of working days in a month;

n h – number of working hours per day;

η – hole utilization factor (BUR).

Borehole utilization rate. The hole utilization ratio is the ratio of the used hole depth to the original depth. When explosive charges explode in boreholes, the rock is not torn off to the entire depth of the boreholes; part of the borehole depth is not used and remains in the hearth mass, which is usually called a glass.

To determine the CSI 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 the charges, you need to measure the depth of all the glasses and determine the average depth of the glass, from which you can find the average value of the KIS. Therefore, in order to determine the average value of the CIS, it is necessary to divide the average advance of the face by the average depth of the hole.

where: l з – length of the hole charge;

l w – hole depth.

If the face advance per cycle is specified, then the average hole depth can be determined by dividing the face advance per cycle by the average value of the hole.

The magnitude of the CIS depends on the strength, fracturing and layering of the rocks to be blasted, the area of ​​the face, the number of open surfaces in the blasted mass, the operability of the explosive, the depth of the holes, the quality of the hole stopping, the order of blasting of the charges and other factors. With the correct determination of all parameters and strict implementation of the technology for conducting blasting operations, the value of the KIS should be no less than the following values.

Table 4.4.

Table 4.5

Numerical values ​​of the indicator γ

 vv, kg/m 3
, units	1.843	1.892	1.940	1.987	2.033	2.125	2.214	2.301

 explosive - volumetric weight of explosive in charge, kg/m 3

The distance between the contour charges is determined by the formula (m):

(4.6)

Where: K 0- numerical coefficient taking into account the interaction of adjacent contour charges and energy losses due to the expansion of detonation products in the volume of the hole, units;

L zk- length of cutting of contour holes (determined from the table), m;

L to- length of contour holes, m.

Table 4.6

Numerical coefficient value K 0

Table 4.7

Reduced length of driving of contour charges L sq / S vyr

Coefficient	Linear density of loading contour holes P to, kg/m
rock strength	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 coefficient of proximity of contour holes is determined by the formula:

(4.7)

At  cc= 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- width of excavation at soil level, m

The area of ​​the part of the face that falls on the contour row is equal (m2):

(4.11)

To improve the quality of rock development at the level of the end parts of contour holes, an additional charge weighing (kg) should be placed in the bottom of the latter:

The amount of explosives per contour break is determined by the formula (kg):

During preliminary contouring development, the specific consumption of explosives is determined taking into account the depth of work N(m) according to the formula (kg/m 3):

(4.14)

It should be borne in mind that when reducing the depth of work, the value q to should not be less than the value determined by formula (4.3).

The distance between contour holes is calculated using formula (4.6), while the value L zk determined according to table (4.8).

Table 4.8

Reduced length of contour charges during preliminary delineation of excavation

Coefficient	Depth of work N, m
fortresses	less than 100	100-200	200-400	400-600
breeds, f	Linear loading density of contour holes 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 at the bottom of contour holes is determined by the formula (kg):

Number of contour holes N to and consumption of explosives for delineation of mine workings Q to calculated using formulas (4.10) and (4.13)

After determining the parameters of contour blasting, we move on to calculating the parameters of loading and placing cutter and breaker holes. The basis of the calculation is the value of the specific consumption of explosives for crushing rock within the drilled volume.

During subsequent delineation, the core of the face is broken under conditions of stress in the surrounding rock mass, which leads to the need to increase energy costs for crushing rock in the drilled rock mass. In this case, you must first determine the characteristic value of the length of the breaker holes, taking into account the degree of such influence (m):

(4.16)

Depending on the actual length of the blast holes L, 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 using the formulas (kg/m 3):

At L out  L  :

(4.17)

At L off  L  :

(4.18)

Where: e bb- conversion factor taking into account the type and density of the explosive used.

Table 4.9

The value of the coefficients e bb

During preliminary delineation of a mine working, the main rock volume is broken under conditions of partial unloading, which allows, given the length of the breaker holes, L off  L  reduce the specific consumption of explosives to the value determined by formula (4.17)

After determining the specific consumption of explosives, the parameters for placing holes in a straight cut are calculated. The value of the specific consumption of explosives in the cutting is determined taking into account the overall efficiency of rock breaking in the working face:

(4.19)

Where: N time- number of cutting holes, units;

P vr- linear density of their loading, kg/m;

L vr- length of cutting holes, m;

L zb- stope length, m.

Absolute value L zb determined from the tables below, followed by division by e cc, which makes it possible to take into account the type of explosive used.

Table 4.10

during subsequent delineation of the mine workings

Coefficient	Depth of work N, 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 driving holes during preliminary delineation of mine workings

Rock strength coefficient	L zb / S vyr
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 drilling area of ​​the cut is determined by the formula (m2):

(4.20)

The amount of explosives in the cut is determined by the formula (kg)

(4.21)

Since in straight cuts the rock is crushed under conditions of one free surface, to facilitate the work of the cutting 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: Wmin- distance from the well to the nearest cutting hole working for this well, m;

d sp- cutter hole diameter, m.

Knowing the area of ​​the cut and taking the cross-sectional shape in the form of one or another flat geometric figure, you can determine the dimensions of the cut section and the parameters for placing cut holes (Fig. 4.3):

Square:

Slotted:

(4.27)

(4.28)

Fig 4.3 Examples of placement of boreholes in straight cuts.

After calculating the cutting parameters, we proceed to calculating the breaking parameters.

The total number of break holes (including soil holes) is determined by the formulas (pcs.):

During subsequent contouring:

(4.30)

For preliminary contouring:

(4.31)

Where: R otb- linear density of loading blast holes, kg/m;

e otb, e k- conversion coefficients for fender and contour charges, respectively.

The distance between soil holes is calculated by the formula (m):

(4.32)

The line of least resistance of soil holes is determined by the formula (m):

(4.33)

The number of soil holes and the area of ​​the part of the face that falls on these holes is determined by the formulas:

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

(4.35)

The approximate size of the fencing hole drilling grid is determined by the formula (m):

(4.36)

During preliminary delineation of the workings S to = 0.

The amount of explosives for rock removal within the core and soil zones is determined by the formula (kg):

Based on the calculations and the hole location diagram, a summary table of blasting parameters by shape is compiled.

Drilling and blasting parameters table

Rice. 4.4 Drill hole layout.

a – drill hole layout; b – charge design; 1 – explosive cartridge;
2 – electric detonator.

After calculating all the parameters of the drilling and blasting complex, a drilling and blasting passport is drawn up.

The drilling and blasting blasting certificate must contain a diagram of the location of the blast holes (in three projections), indicate the number and diameter of the holes, their depth and angles of inclination, the number of blasting series, the blasting sequence, the size of the charges in the blast holes, the total and specific consumption of explosives, the consumption of detonators, the length of the internal stopping of each hole and the total amount of stopping material for all holes, as well as the time of ventilation of the hole.

To explain the text part of this section, the note should provide the corresponding diagrams (hole layout diagram, charge design diagram in the hole, cutting diagram, connection diagram of detonators in the explosive network).

Calculation of an electrical explosion network.

When electrically exploding charges, it is possible to use all known schemes for connecting resistances into a circuit. The choice of EM connection scheme depends on the number of EDs to be exploded 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 maximum value of the circuit resistance specified in the device passport. When using power and lighting lines, the resistance of the explosive circuit is determined, then the amount of current passing through a separate ED is calculated, and this value is compared with the guaranteed current value for a failure-free explosion. For the guarantee current it is accepted - for 100 ED equal to 1.0 A, and when exploding ED in large groups (up to 300 pcs) 1.3 A and at least 2.5 A when exploding with alternating current.

In a series connection, the ends of the wires of adjacent electric motors are connected in series, and the outer wires of the first and last electric motors are connected to the main wires going to the current source.

The total resistance of the explosive circuit with a series connection of the ED is determined by the formula:

, Ohm (4.38)

Where: R 1- resistance of the main wire in the area 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 end wires of the ED to each other and to the main wire, Ohm;

n 1- number of EDs connected in series, pcs;

R 3- resistance of one electric motor with end wires, Ohm.

The resistance of the wires is determined by the formula:

Where: ρ – resistivity of the conductor material, (Ohm*mm 2)/m;

l– conductor length, m;

S– conductor cross-section, mm 2.

When carrying out blasting operations, wires for industrial blasting operations of the VP grade with copper conductors in polyethylene insulation are used as connecting wires and for laying temporary blasting lines. The wire is produced as single-core VP1 and two-core VP2x0.7.

Cables of the NGShM brand are intended for laying permanent blast lines. Current-carrying conductors are made of copper wire. The insulation of current-carrying conductors is made of self-extinguishing polyethylene.

In exceptional cases, in agreement with the Gosgortekhnadzor authorities, VP2x0.7 wire may be used as permanent blast lines

Table. 4.12

Table. 4.13

Table 4.14

Drilling holes

Hole drilling is done with hand drills, hammer drills, and drilling rigs.

Hand drills- used for drilling holes up to 3 m deep in rock with f  6. Drilling is done directly from hands or from light supporting devices (SER-19M, ER14D-2M, ER18D-2M, ERP18D-2M). Electric core drills are used when drilling through rock with f  10 (SEK-1, EBK, EBG, EBGP-1).

Where: n- number of drilling machines;

k n - machine reliability coefficient (0.9);

k 0- coefficient of simultaneity of machine operation (0.8 - 0.9).

The number of drilling machines is determined on the basis of 4 - 5 m2 of face area per drilling machine.

Hammers- used for drilling holes in rocks with f  5 (PP36V, PP54V, PP54VB, PP63V, PK-3, PK-9, PK-50).

Drilling productivity is determined by the formula (m/h):

(4.45)

Where: k d- coefficient depending on the diameter of the hole (0.7 - 0.72 for d w = 45 mm; 1 for d w = 32 - 36 mm);

k p- coefficient taking into account the type of hammer drill (1.1 for PP63V, PP54; 1 for PP36V);

A– coefficient that takes into account changes in drilling speed in different rocks (0.02 at f = 5-10; 0.3 at f = 10-16).

Drilling rigs. Hole drilling is carried out using drilling rigs or mounted drilling equipment mounted on loading machines.

The selection of a drilling rig for drilling holes in a horizontal mine is made taking into account the following factors:

The type of drilling machine must correspond to the strength of the rocks in the face being drilled;

The dimensions of the drilling zone must be greater than or equal to the height and width of the face being drilled;

The maximum length of the drilled holes according to the technical characteristics of the drilling machine (installation) must be consistent with the maximum length of the holes (according to the drilling and drilling drilling passport);

The width of the drilling rig should not be greater than the vehicles used.

Drilling productivity is determined by the formula (m/h):

(4.46)

Where: n- number of drilling machines on the installation, pcs;

k 0- coefficient of simultaneity in the operation of machines (0.9 - 1);

k n- installation reliability coefficient (0.8 - 0.9);

t- duration of auxiliary work (1 - 1.4 min/m);

v m- mechanical drilling speed (m/min).

Table 4.5

Drilling speed

Duration of drilling holes (hours):

Where: t p– preparatory and final work (0.5–0.7 hours).

Ventilation design.

The design of underground mine ventilation is carried out in the following sequence:

1. The ventilation method is selected;

2. The pipeline is selected and its aerodynamic characteristics are determined;

3. The amount of air required to ventilate the workings is calculated;

4. A local ventilation fan is selected.

The installation location of the local ventilation fan (VMP) and the direction of the ventilation pipeline 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 streams, and safety zones.

Ventilation methods.

Ventilation of workings is carried out using injection, suction or combined methods.

With the injection method, fresh air is supplied through pipes to the face, and contaminated air is removed through the workings. The main advantage of this method is the effective ventilation of the bottom-hole space when the ventilation pipes lag significantly behind the face chest. In this case, it is possible to use flexible pipes. However, due to the fact that gases are removed over the entire cross-section and along the length of the mine, gas contamination occurs, which leads to the need to install fans of higher capacity and pressure and lay air ducts with pipes of larger diameter. This method is the most widespread.

With the suction method, toxic gases do not spread through the workings, but are sucked out through the pipeline, and fresh air enters the bottomhole space along the workings. The main advantage of this method is that with a sufficiently small distance of the end of the pipeline from the face chest, not exceeding the suction zone, the working face is ventilated much faster than with other methods, and gas contamination of the main part of the working face does not occur. This method can be used to ventilate workings when the main sources of production hazards are concentrated in the bottomhole zone. The insertion method cannot be used when driving workings through gas-bearing rocks, when operating rolling stock with internal combustion engines in them, or when there are other sources of hazardous emissions distributed along the length of the workings.

The combined method involves the use of two fans, one of which works for exhaust, and the other, installed near the bottom, for injection. This method of ventilation combines the advantages of injection and suction methods. In terms of ventilation time, this method is the most effective. The disadvantages of this method is the cluttering of the mine with ventilation equipment.

Rice. 5.1 Ventilation schemes for dead-end workings.

a – injection; b – suction.

1 – fan; 2 – pipeline.

Table 5.1

Coefficient value R 100

Pipe diameter,	Metal	Type M	Textovinite
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 resistance of the pipeline. The pressure created by the fan during its operation on the ventilation pipeline is spent on overcoming friction and local resistance, as well as on the speed pressure when air leaves the pipeline or enters it, during suction ventilation.

The aerodynamic friction resistance of the pipeline is determined by the formula:

, N*s 2 / m 8 (5.2)

Local resistance of ventilation pipelines is usually created by elbows, tees, branches and other fittings of pipes. The local resistance values ​​are given below.

Table 5.2

Resistance (N*s 2 / m 8

The dimensions of the cross-section of horizontal mine workings in the open air depend on its purpose and are determined based on the dimensions of the rolling stock and the equipment located in the working, ensuring the passage of the required amount of air, 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 excavation of a rectangular-vaulted shape with roof anchors.

Rectangular-vaulted sections are used when excavating workings without support or with the construction of lightweight support structures. The height of the vault in sections from 2 to 6.8 m2 is?. working width.

The clear cross-sectional area is the area along the internal contour of the support installed in the excavation

Calculation of the mine cross-section

Cutting width

b=b c +2c= 0.95+2 0.3=1.55m

where b c is the width of the scraper, m;

c is the gap between the scraper and the side of the excavation, m.

In the development of the type under consideration, people are allowed to walk only when the scraper installation is not working. Thus, the height of the excavation in the open air is assumed to be minimal, i.e. 1.8 m.

Arch height

Height of the cut on the side (up to the heel of the arch):

1.8 - minimum production height according to PB

Based on the calculated clear cross-sectional area, the nearest larger standard cross-section from the table is taken. 2 (Tutorial “Conducting horizontal exploration workings and chambers” Authors V.I. Despite, V.A. Kosyanov Moscow 2001).

The standard cross-section of the PS excavation is accepted - 2.7

The main dimensions of the excavation cross-section in the clear:

Working width, mm - b = 1550 mm

The height of the excavation to the heel of the arch, mm - h b = 1320 mm

Working height, mm - h = 1850 mm

Radius of the axial arch of the arch, mm - R = 1070 mm

Radius of the side arch of the arch, mm - r = 410 mm

Clear cross-sectional area of ​​the excavation, m2 - Sv = 2.7 m2.

For excavations with roof anchors:

where is the height of the excavation along the side, taking into account the exit of the anchors along the roof into the excavation by the value d = 0.05 m.

Calculation of strong dimensions of the support, drawing up a fastening passport

Due to the small cross-section of the excavation, short service life, mining and geological conditions and available materials, we use metal expansion anchor support AR-1

All calculations of the strength of anchorage in the hole were made using formulas from the reference book “Theory and practice of using anchorage” Author A.P. Shirokov. Moscow "Nedra" 1981

ts - rock friction angle, 30 degrees

D - diameter of spacer coupling, 32cm

h - spacer sleeve height, 30cm

y szh - rock compressive strength

b - half the angle of a symmetrical wedge, 2 degrees

p 1 - angle of friction between steel and steel, 0.2 degrees

The required length of the anchor L a in the roof and the height of possible rock fall out of the excavation is found from the expressions:

L a = b+ L 2 + L 3 =0.04+0.35+0.05=0.44m;

where L 2 is the depth of the anchors beyond the contour of a possible rock fall (taken equal to 0.35 m); L 3 - length of the anchor protruding beyond the excavation contour, L k = 0.05 m; a n = half-span of the excavation in the tunnel, m; h is the height of the opening in the tunnel, m.

Coefficient characterizing the stability of the excavation sides;

Coefficient characterizing the inclination of the sliding prism in the sides of the excavation (taken according to Table 12.1. Theory and practice of using roof bolts. Author A.P. Shirokov. Moscow "Nedra" 1981);

ц b - angle of internal friction (resistance) of rocks in the sides of the excavation; K k - coefficient taking into account the decrease in the strength of rocks in the roof of the mine (accepted according to Table 13.1);

f k - coefficient of rock strength in the roof of the workings;

K szh is the coefficient of concentration of compressive stresses on the excavation contour, the value of which is taken from the table. 12.2;

g - average specific gravity of rock strata overlying the excavation to the surface, MN/m 3 ; H - excavation depth from the surface, m;

K b - coefficient taking into account the decrease in the strength of rocks in the sides of the excavation, the value of which is taken according to Table 12.1;

f b - rock strength coefficient according to M.M. Protodyakonov in the sides of the mine.

We accept the length of the anchor in the roof L k = 0.5 m.

Due to the fact that w0, the sides of the excavation are not anchored.

Roof area supported by one anchor

where F k is the roof area supported by one anchor, m 2 ;

P k - strength of anchor fastening in a hole drilled in the roof;

The safety factor, taking into account the uneven distribution of load on the anchor and the possibility of additional loading from overlying layers, is taken equal to 4.5;

b - excavation inclination angle, degree 0 0

Distance between anchors in a row:

where L n is the step of installation of anchors along the width of the excavation, m;

L y - distance between rows of anchors, m, assumed to be 1.4 m

Number of anchors in a row

where L b =1.33b=1.331.55=2.06m is the part of the excavation perimeter that is subject to roof anchoring, m. Where b is the rough width of the excavation.

Accepts 2 anchors in a row.

Drawing up a fastening passport.

Clear cut width:

B = B + 2m = 950 + 3002 = 1550mm.

Height of the vault

h o = b/3 = 1550/3 = 520mm.

Rough cut height

h 2 = h + h o + t = 1320 + 520 + 50 = 1890mm.

Rough cutting wall height

h 3 = h + t = 1320 + 50 = 1370mm.

Radius of the axial arc of the arch of the dissection

R =0.692b = 0.6921550?1070mm.

Radius of the lateral arc of the arch of the dissection

r = 0.692b = 0.6921550?410mm.

Clear cross-sectional area:

S St = b(h + 0.26b) = 1.55(1.32 + 0.261.55)? 2.7m2

Perimeter of the cross-section of the cut in the clear:

P = 2h + 1.33b = 21.32 + 1.331.55 = 4.7 m.

Cross-sectional area of ​​the rough cut:

S hf = b (h 3 + 0.26b) = 1.55 (1.37 + 0.261.55) = 2.75 m 2.

Perimeter of the cross-section of the rough cut:

P = 2h + 1.33b = 21.37 + 1.331.55 = 4.8m

Distance between anchors in a row: b 1 = 1200mm.

Distance between rows of anchors: L = 1.4 m

Depth of holes for anchors: l = 500mm.

Diameter of holes for anchors: = 43mm.

The maximum lag of the anchor support from the face chest is assumed to be 3 m.

Scheme for calculating cross-sectional dimensions when using scraper equipment in the development of a rectangular-vaulted sectional shape.

1. Selecting the shape and calculating the dimensions of the cross-section of the excavation

When carrying out workings, two types of mining operations are distinguished: main and auxiliary.

The main mining operations are those that are carried out in the face of the workings and relate directly to the excavation and securing of the workings.

Auxiliary operations are those that provide normal conditions for performing the main tunneling operations.

The cross-sectional area of ​​the excavation depends on the purpose and dimensions of the equipment located in it. There are cross-sectional areas of horizontal workings in the open, rough and after excavation. The clear area is determined by the dimensions of the excavation up to the support, minus the areas occupied by the ballast layer and ladder in the excavation cross-section. The rough area is the designed area in the excavation. When determining this area, the area occupied by the support, ballast layer, ladder and tie (with frame supports installed staggered) is added to the clear area. The actual area that is obtained as a result of excavation is usually 3-5% or more greater than the designed area.

The cross-sectional dimensions (width and height) of haulage workings depend on the overall dimensions of haulage cars and electric locomotives, on the rail tracks, the method of movement of workers through the workings and the amount of air supplied for ventilation.

If there are rail tracks in the excavation for the movement of people, a path (passage) with a width of at least 700 mm is provided, which must be maintained at a height of 1800 mm from the level of the ladder (ballast layer).

Based on specific conditions: f =16; stability - average; service life of the excavation is 16 years, we choose a vaulted excavation shape, sprayed with concrete fastening

1. Calculate the cross-section of the excavation height.

a. Height of the rail track structure h 0, mm

h 0 = h b + h w + h p + h p, mm;

Where: h 0 - the height of the upper structure of the excavation path, is selected according to the standards stipulating ESP, mm;

h b - height of the ballast layer, mm;

h p - height of the lining under the rail, mm;

h r - height of the rail track, mm;

h 0 = 100 + 420 + 20 + 135 = 375 (mm).

2. Height of rolling stock h, mm

3. Height of the straight-walled section of the excavation.

h 1 = 1800 (mm).

4. Clear height of excavation.

h 2 = h 1 +h b +1/3h w, mm;

h 2 =1800+135+20+1/3*120=1995 (mm).

Where: h 1 - height of the straight-walled section of the excavation, mm;

h b - height of the ballast layer, selected according to the standards stipulating ESP, mm;

h w - height of the sleeper beam, mm;

5. The height of the workings in the blacker.

h 3 = h 0 + h 1, mm;

h 3 =375+1800=2175 (mm).

6. Clear height of the vaulted ceiling.

h h =1/3*V, mm;

h h =1/3*2250=750 (mm).

7. The height of the vaulted ceiling in the basement.

h 5 = h h + T cr. , mm;

h 5 =750+50=800 (mm).

8. The clear width of the excavation is calculated.

B= n+A+m, mm;

B=200+1350+700=2250 (mm).

Where: B is the clear width of the excavation, mm;

n is the gap between the support and the rolling stock, mm;

A is the width of the rolling stock, mm;

m - free passage for people, mm;

9. Width of working in the black.

B 1 =B+2* T cr. , mm;

B 1 =2250+100=2350 (mm).

10. Clear cross-sectional area.

S St. = V*(h 2 +0.26*V)

S St. = 2250*(2745+0.26*2250) =7.4 m2

11. Cross-sectional area in black.

S black = V 1 *(h 3 +0.26* V 1)

S black = 2350*(2960+0.26*2350) =8.3 m2

12. Air flow speed.

V = Q air / S c in, m/s;

V = 18/7.4 =2.4 m/s;

Where: V is the speed of movement of the ventilation stream through the mine working, regulated by safety rules, m/s;

Q air - the amount of air passing through the workings, m 3 /s;

S c in - cross-sectional area of ​​the excavation in the open, m 2 ;

Since V = 2.4 m/s, then 0.25? V? 8.0 satisfies the requirements of the HPB, therefore, this section was calculated correctly.

13. Section in the penetration.

S pr =1.03* S black, m

S pr =1.03* 8.3 =8.7 (m)

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