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CHAPTER 4

SURFACE MINING METHODS (PLACER MINING)

4.1 Alluvial deposits (or) Placers

Placer ids a term applied to deposits of ore or more minerals that have accumulated inquantities of economics importance through the natural processes of weather and

concentration. In general, the heavy and not easily decomposed minerals in the parent rock

are freed as the rock is broken down and are transported, sorted and collected into workable

deposits by the action of water. Some placers, however, are residual placers formed by the

weathering and removal of the parent rock, the valuable minerals being left behind.

Eluvial is a term applied to the placers that are found close to the placers that are

found close to the parent rock. Thus free gold found in the ground on a hillside below the

outcrop of a gold-bearing vein is eluvial. Alluvial is the broad term applied to placers formed

by the mechanical action of moving water of streams, lakes or oceans.

The specific gravity of the more important placer minerals is listed below, the

variations being due to the presence of impurities:

Gold 15.5 to 19.3

Platinum 14 to 22

Cassiterite(tin) 6.6 to 7.1

Diamond 3.2 to 3.52

` Garnet 3.15 to 4.3

Monazite 4.9 to 5.3

Magnetite 5.16 to 5.18

4.2 Classification of placer deposit

Various classifications of alluvial deposits are following:

(i) Residual or eluvial deposits(ii) River and stream deposits(iii) River terrace or bench deposits(iv) Bench or marine placers(v) Deep leads or buried placers
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4.2.1. Residual or eluvial deposits

Residual or eluvial deposits, including the so-called hillside occurrences, which are

formed by the weathering of out-crops, and usually located on the gentle slopes of valleys.

4.2.2 River and stream deposits

River and stream deposits, including gulch and creek deposits, and river bar deposits,

which are formed in or bordering rivers and streams.

The weathering of auriferous reefs, which are often of quartz, results in the latter being

broken up and carried down the valleys, as sand and gravel, heavy minerals occurring in the

reefs accompany the gravel, with the result that native gold, magnetite and heavy black sands,

etc. are deposited in the valleys and drainage channels and may eventually form payable

alluvial deposits of gold.

Similarly, cassiterite is usually associated with granite intrusions, occurring there in as

veins, stringers, stock works and the like: payable tin alluvial may be derived from the

weathering of such occurrences in the granite masses.

4.2.3 River terrace or bench deposits

River terrace or bench deposits, which occur on the flanks of valleys and are the

remnants of old river deposits, formed by the river cutting itself a newer and deeper channel

in the bed-rock; rivers and streams may change their courses and levels from time to time,

with the result that alluvial deposits may occur not only in the present river beds, but also in

older buried channels, which form benches on the slopes of the existing valleys.

4.2.4 Beach or marine placers

Beach or marine placers which occur along coastal strips in many countries, having

been formed by the sorting action of waves, which tends to concentrate beach material, that

has resulted from the erosion of rocks, together with any heavy minerals that may have

resisted weathering.

Such deposits are represented by the black sand beach deposits, containing limonite,

rutile, monazite, zircon, etc. found along the coasts of Southern India, Australia, U.S.A,

Brazil, etc. and the diamonds come to be found in river beds and in the beach sands along


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some coastal areas, considering they are of volcanic origin; it is possible that alluvial deposits

originally derived their diamonds from the erosion of the upper portions of kimberlitic popes,

and so transported to the sea, where wave action re-deposited the diamonds along the shore.

4.2.5 Deep leads or buried placers

Deep leads or buried placers, which are ancient deposits that have been buried under

overburden, varying in thickness from 100 to 1000 ft and more.

4.3 Economic placer minerals

Placer deposits were formed as a result of destruction of primary deposits by process

of rock weathering. Placers are deposits of sand, gravel, or other alluvium, containing

particles of valuable minerals in workable amounts. Native gold is the most important placer

mineral; a large part of the world's output of platinum and cassiterite is derived from gravels,

other minerals for which alluvial deposits are regularly worked include monazite, columbite,

ilmenite, zircon, diamond, sapphire, ruby and other gems occur occasionally in gravels.

4.3.1 Gold

Gold is the only metal which pure and in mass has a yellow color. Small amounts of

silver reduce the depth of color, while copper darkness the tint; the redder tint of the British

sovereign is due to inclusion of 2 parts of copper.

Gold at all temperatures possesses a far greater degree of malleability and ductility

than any other metal. It is softer than silver but harder than tin and has a hardness of 2.5

according to Moh's Scale. The specific gravity of gold is about 19.3 when pure, but may vary

from 15 to 19.

The gold in placer deposits is derived from the weathering and breaking down of gold-

bearing veins to form eluvial deposits near the source of origin and alluvial deposits in which

the material has been transported by hydraulic action and deposited at lower levels. The best

conditions for concentration of gold in auriferous gravels are when the river gradient is mode-

rate, under balanced conditions of erosion and deposition.

4.3.2 Platinum
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Malleable and extremely ductile in its pure state, the addition of a very small quantity

of iridium renders it one of the hardest metals known. Platinum is the most common member

of a group of metals, which includes palladium, osmium, iridium, ruthenium and rhodium, all

of which are usually found in intimate association. Except in rare instances native platinum is

invariably alloyed with 10% to 40% of varying amounts of the other metals of this group and

gold and iron.

4.3.3Tin

Cassiterite or oxide of tin is the only mineral of this metal of economic importance.

The two most important modes in which it occurs are as lode and alluvial.

Lode tin is found together with other minerals in veins formed by the filling of cracks

or fissures in the country rock, generally by quartz together with tourmaline and mica.

Alluvial tin occurs in loose deposits of a sandy clay nature laid down by water and

generally forming flat or gently undulating lowland country. The tin mineral in these deposits

has been derived from the erosion of softer rocks in which the stanniferous lodes, stock works

or impregnations occurred.

Tin ore is always associated with granites rocks and rocks of analogous composition

rich in quartz. It also occurs in lodes or stringers traversing schist or shale near their contact

with granite. Cassiterite contains approximately 78.5% of metallic tin, has a hardness of from

6 to 7 according to Moh's scale of hardness and a specific gravity of 6.8 to 7.2.

4.3.4 Tungsten

The most important sources of tungsten are wolfram, (FeMn)WO4 , and scheelite,

CaWO4. Tungsten is used as ferro-tungsten or tungsten powder for the production of high-

speed steels, as powder for making lamp or valve filaments and as the carbide for cutting tools

and dies, mining equipment and armour-piercing projectiles

Wolfram is often associated with cassiterite in quartz and pegmatite veins traversing

granitic rocks and sedimentaries in contact with such rocks. It is less stable than cassiterite

and on account of its good cleavage is easily weathered.

Tungsten deposits have been formed under widely different conditions of temperature

and pressure, but they characteristically occur in or near the more acid igneous rocks. Most of


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the commercial primary occurrences are in veins or replacement bodies which are near

igneous intrusions.

Thus the wolfram placers of China have been the world's principal source of tungsten.

In Southern Myanmar alluvial wolfram is recovered as a by-product from tin concentrates.

4.3.5 Diamond

If concentrates, suspected of containing diamonds, are turned over, under water,

slowly and carefully, any existing diamonds will be easily distinguished by the brightness of

the stones, as compared with other existing crystals, due to the very high refractive index of

the diamond.

4.4 Classification of Placer Mining Method

Placer mining is essentially an exercise in large scale earth moving using plant and

equipment that is also employed in such other undertakings as land reclamation, the dredging

of harbors and waterways, road construction and quarrying. The basic systems and methods

used are also the same, only the objectives differ, and there is a need for greater precision

when mining along the boundaries of deposits and in cleaning up at bedrock. The valuable

constituents in a placer are rarely distributed evenly and, while the bulk of the material

(overburden) can be removed as in any other undertaking, the pay materials must be taken up

separately and fed to the treatment plant at a required rate and in a designated form.

As a result, some aspects of placer mining economics are concerned more with

selectively than with the actual quantities moved per unit of time and, since the ultimate

purpose of mining is to make a profit, maximum benefits accure to operators who plan every

aspect of an p\operation in detail while still retaining sufficient flexibility of mind and the

ingenuity to deal with the unexpected. Vast sums of money have been spent on improving

machine design and capabilities and the miner has an extensive array of models to choose

from, the task is to select the best of the alternatives for each particular duty. Governing

factors are concerned with how well a particular machine or combination of machines is

likely to adjust to conditions that may in some ways be different to those for which they were


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designed. Other aspects of increasing importance relate to the environmental impact of

mining.

Therefore, Placer mining methods can be classified as follows:-

I. Hand miningII. Mechanical mining(A) Dry Mining system

(i)Opencast mining with power shovels

(ii)Opencast mining with draglines

(iii)Opencast mining with scrapers and loaders.

(B) Wet Mining System(i)Hydraulic mining

(Including water supply, monitors, hydraulic elevators, sluice boxes,

gravel pump etc. )

(ii)Dredging

4.5 Choice of mining method

The first consideration is whether to mine wet or dry. In some cases there may be no

reasonable alternative to a system that is wholly wet (eg. offshoe dredging), or wholly dry (eg.

desert mining). At other times there is a choice and some methods may incorporate elements

of both system (eg. dry mining followed by dredging or hydraulic sluicing.

Each placer has some distinctive features and, hence, each requires separate study to

determine by what method it can be most economically mined. Basic factors influencing the

choice are mainly:

(a)deposit size and value

(b)sediments and minerals characteristics

(c)availability of water

(d)bedrock characteristics and geometry and

(e)the total environments


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Table: Dry versus Wet mining system

Dry mining Wet mining

Application

Systems built around

Controlling factorsfor selection

Advantages

Disadvantages

Shallow surface deposits, tightly

compacted or indurate sands,

irregular geometry, high level

dunes, desert environment.

Bulldozers, scraper loaders,

articulated front end loaders,

dragline, hydraulic excavators.

Proposed scale of miningminerals distribution and value,

location and physical

characteristics, slope and texture

of mining floor surface and bed

rock geometry, availability of

water.

Ability to handle group of small

deposits, constant feed rate

under widely differing mining

conditions, selective mining

leads to optimization of feed

grade control, recoveries may

approximate 100%.

High unit operating costs

difficulties in handling

occasional large volumes of

water, requires firm base forvehicle movement, needs large

on-site workshop facilities and

stock of spare parts.

All environment where ample water is

available for mining and treatment

including shallow surface deposits,

high-level dunes, marine deposits.

Pumps and monitors, suction and

bucket dredgers, bucket wheel

dredgers, clamshell dredgers, jet-lift

dredgers, hydraulic excavators.

Proposed scale of mining, deposit sizeand grade, location and physical

characteristics, slope and texture of

mining floor. Bedrock geometry,

amount of water, position of water

table.

Mining and processing can be

incorporated in one unit, low unit

mining costs, closer supervision and

control, only possible method in

excess water conditions.

Mining losses sometimes high, less

selectivity in mining, possible high

relocation costs, large water

requirements, ecological problemsmay affect large sections of

environment.


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4.5.1Hand mining

4.5.1.1 Manual method

Manual methods of placer operation are restricted to pioneer conditions, small rich

deposits, thin scattered deposits, and remote places without roads or convenient access, or

where it is impracticable to obtain machinery or to develop water supplies for hydraulic

mining.

Hand method are used widely throughout the world for mining small deposits in

regions that are largely inaccessible to machinery and for cleaning up pockets of gravels

remaining untouched from previous operations. Whole communities may be involved in hand

mining operations in large areas of provenance where innumerable small shallow deposits

occur and labor intensiveness is a major consideration.

Work output varies very much with the climate. A good strong labor in a temperate to

cold climate can be expected to shovel between 6m3and 8m

3of loose gravels per day in some

tropical areas. Various types of sluices, rockers and pens used for recovering the valuable

minerals have capacities that differ according to the case with which the particles are liberated

from their matrices of clay, sand and gravel pans used for concentrating valuable heavy

minerals by hand vary from community to community.

There are two fundamental methods of hand operation, shoveling into sluice boxes and

ground sluicing.

4.5.1.2 Ground sluicing

Ground sluicing methods have been developed independently in many parts of the

world over a long period of time dating back to antiquity. The usual method is to construct a

dam across a watercourse above the area to be mined and to channel a stream of water along

flumes leading to the pay gravels. This water is directed into a cutting in the gravels and

material shoveled in from the sides is broken up and slurred manually. The valuable particles,

generally of gold are caught up in natural riffles formed by the unevenness of the bed or

behind artificial riffles in an upstream direction and the method is reasonably successful

where steep slopes are available for tailings disposal.

Requirements for ground sluicing:

(a)Shallow gravel, rarely more than 6 to 8 ft deep.


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(b)Sufficient grade for the available water to carry the loosened sluiced muck on

grades of 17 to 25 ft per mile: grades for gravel are much steeper.

(c)Plentiful water. Here again data are lacking. Long ridge states that it takes about 6

times as much water to move material in a ground sluice as to do the same work in a box

sluice. Scanty water supply may sometimes be supplemented by pumping back from settling

pools at end of sluice.

(d)Dump room for tailings at lowers end of sluice may be provided by natural grade of

bedrock or surface: drag scrapers may move tailings from end of sluice.

Hand miners have learned, from experience, to control flume velocities below the

critical erosion velocities for the particular materials over which they flow. Ground textures

differ widely and wooden flumes are sometimes used when steep gradients cannot be avoided.

4.5..1.3 Instruments used in hand mining

The pan and rocker are used by prospectors in searching for placers, by miners for

washing gravel on a small scale and by engineers for recovering gold from samples

obtained in placer examinations. For testing placer deposits and working pockets and

small placer deposits, the pan, the rocker and the long tom are used.

Pan and Bates

The pan is a circular steel dish from 10 to 16 in. in diameter at the top, from 2 to 2.5

in. deep, and with sides sloping at 35deg or 40 deg to the horizontal. A pan of gravel is placed

in water and stirred by hand to break up lumps of clay.

Larger stones are picked out and the pan is given a gyratory motion to settle the

heavier particles. From time to time the pan is tilted and the surface layer of material is

washed off. At the end of the operation a little black sand (if the placer contains black sand) is

left with the gold, which can be collected by adding a little mercury, or the material can be

dried and the sand removed by a magnet. By diligent work an experienced man can pan about

1 cu.yd of gravel in 10 hours.

The bates is a flat, conical woolen or iron pan from 16 to 30 in. in diameter. It is used

in much the same way as the gold pan.


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10 to 16 in.

2.5 in.

40 deg

Fig. Pan

The Rocker

The gold bearing gravel is placed on the screen. The gold and fine sand are washed

through the screen and the stones remaining are cleaned out. The chute directs the material to

the upper end of the bottom, which may be covered with small transverse riffles or canvas.

Canvas or burlap is a better saver of fine gold than riffles. Waste material passes over a

tailpiece at the end of the rocker. In clay ground, the clay should be broken up in a puddling

box ahead of the rocker. Sample can conveniently be carried to the rocker in pails.

Fig. Cross section of a rocker

Rocks range in length from 6 to 12 ft. and in bottom width from 14 to 20 in. Long

rockers are better savers of fine gold. Holes in the screens and from to in. in diameter.

The slope of the rocker should be adjusted to the nature of the gravel and is commonly 1 in


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12, ranging from 1 in 8 to 1 in 20. Two men with a rocker can handle from 3 to 5 cu.yd. of

gravel in place in 10 hours if the ground is easily rocked.

4.6 Hydraulic mining of placers

By Hydraulicking (also called ground sluicing) is meant rending the ground (sands,

peat) from its solid mass in place by a strong water jet with subsequent transport of the

dislodged material by a stream of water to the site of its treatment or storage.

In hydraulic placering washing of sands in wet jiggers and removal of tailings

thereafter are also affected by means of a strong water flow. In hydraulicking water is

supplied along pipes to a hydraulic giant (monitor). The water escapes from the nozzles of the

monitor at a velocity of up to 20-25 m/sec. In the absence of a valley slope needed to carry

away the sands hydraulic elevators or pump dredges are used for lifting the sands to the

required height.

The basic conditions for the application of hydraulic mining of placer deposits are

abundance of water, which is consumed in amounts of 8-60 cu-m per 1 cu-m of mined

ground, the possibility of building up a natural or artificial water head of 2-18 at m, the

possibility of unhindered removal of the washed-out product tailings.

In most cases earth-fill dams built up of sandy-clayey ground are used. Small dams are

made of timber logs and bunched fagots. An opening is provided in dams for an outlet pipe

with a gate to regulate the discharge of water and another pipe to drain off excess water.

Water from storage reservoirs, or from the point where it comes from the river, is delivered to

the consumer along ditches, wooden flumes, troughs or pipelines.

4.6.1 System of hydraulic mining

There are classified according to the way in which the giant jet is used in moving

washed out ground from the working face to the main transport facilities. In hydraulicking of

placers preference is given to the methods with concomitant mining, with a side face and a

fan-shaped face.

4.6.1.1 The fan-shaped-face method

In this method the working ditch is driven across the placer. Hydraulic giants set up in

the ditch wash the ground at the face non-uniformly, tending to accelerate wash out in the

section of the face far from the pit.


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The face progresses in a fan-like fashion, the mud-settling sump in the pit forming the

center of the fan. The angle of the face slew varies from 70deg to 360deg (usually it is 160-

220). As regards unit water consumption and the conditions for its application, the system is

similar to the lateral-face method.

4.6.2 Methods of sand wash

The following methods of sand wash are used:

-with a hydraulic giant per up on the bedrock and with bench undercutting;

-with mechanical loosening of the ground;

-with abandonment of unwashed ground.

The first method finds the widest application. Here the monitor rests of the bedrock.

The bench is first undercut over a distance of 15-30m so as to cause mass caving of the face

top. Then the ground is moved towards the mud-settlong sump of the dredge pump. The

method is practiced in grounds of hardness not above the 3d category and with benches not

exceeding 3m in height. For safety make the hydraulic giant should be put up at a distance

from the face breast not smaller than the height of the bench. After the ground has been

washed off over a distance of 2 to 8m the hydraulic giant is moved on to the working face,

and the ground washing process id repeated.

In washout with mechanical loosening of the ground a bulldozer or a power shovel

first rips the latter. This method is employed when the hardness of the ground exceeds 1.5.

Preliminary ripping of the ground reduces the water consumption in bench undercutting. With

benches up to 5m high bulldozers are commonly used to rip the ground. They pile up the

material near the bench toe where the giant is installed. When the bench is high this is cone by

an excavator, with the pile of loosened ground placed on top of the bench (in the case of a

dragline) or near its to (in the case of a power shovel). This method is more efficient with

stronger rocks and expensive electric power. Water consumption is reduced by 30-50%.

(a)With the use of bulldozer

(b)With the use of power shovel

In the third method the bench is undercut with abandonment of a sand layer on the

bedrock so as to increase the gradient of the ground wash-off surface. The height of the

unwashed ground should be such that the wash-off surface slope is close to 0.02. An increase


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in the gradient considerably reduces the amount of water consumed for the transport of sands

over the major part of the face. The giant jet, aided by a bulldozer, carries the remaining layer

of unwashed ground, 0.5-2m thick, away. Ground washout with part of it left unwashed finds

application in working placers with very slightly sloping bedrock (less than 0.005) and also

when the bedrock has receives which can be penetrated by metal.

4.7 Hydraulic mining

Hydraulic mining or hydraulicing, as it is most commonly called, consists in

excavating gravel banks by the impact of water under pressure, the disintegrated material

being carried by means of water through a line of sluices, where it ie washed for the recovery

of its mineral contents. Prior to mining an alluvial deposit by hydraulicing, it is vital to obtain

information regarding the following points:

(1)the area and depth of the deposit, its average value per cubic yard, and the nature of

its mineral content, to enable an efficient type of plant to be designed, thus avoiding loosed,

The minimum payable value per cubic yard depends entirely on local conditions.

(2)the supply of water available and the head or pressure obtainable, as this forms the

basis for calculating the yardage that can be mined per day, A small amount of water under

high pressure will excavate as much material as a large amount of water under a low head.

(3)the total length of leats, flumes, and pipe-lines required for conveying the water to

the "giant", the working faces.

(4)the nature and depth of depth of bedrock from surface, to determine whether the

tailings can be discharged by means of tailings sluices, or whether elevators must be used, if

the deposit is deep and bedrock is well below the natural drainage level, it may necessitate the

working of the gravel banks in two or more lifts, the bolt gravel being handled by elevators.

(5)the area available for the dumping or disposal of tailings, the construction

impounding dams for tailings may be so expensive, especially in latge scale operations, as to

prohibit hydraulicking.

4.7.1 Hydraulic elevators

Hydraulic elevators as usually employed in connection with hydraulic mining have

their greatest used in deep alluvial deposits, occurring in comparatively flat country, and lying


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below the natural drainage level of the district, where an adequate supply of pressure water

is available, and where dredging is not possible. They are also employed where the grade is

not sufficient to use sluice boxes and to provide dumping room in placer mining, where the

natural fall for the disposal of tailings is lacking. The gravel, broken down by the giants,

together with the spent giant water and seepage is delivered to the bottom of the elevator by

means of ground sluices or bedrock sluices, and is elevated to a sluice at a suitable height

above ground level, so as to obtain the required grade and dumping room.

By means of hydraulic elevators, gravel to a total depth of about 90ft has been

successfully worked, but a depth of about 25ft is the most favorable; the ground to be worked

should be relatively free from big boulders and tree stumps, as these not only hinder work but

may require blasting and breaking up, which adds to the cost of the operation.

The elevator utilizes the velocity of the ascending water, caused by the head or

pressure, to lift or elevate the gravel and water; the water under pressure ascends through the

nozzle (N), and in passing through the throat (T), which has a restricted area, sucks up water

and gravel from the intake or ground section (O), the whole of which is forced up the upraise

or upcast pipe (p) by the force of the jet into the elevated sluice box.

The inclination of the pipe upraise varies from as much as 40deg to 90deg, but is

usually set at an inclination of from 60deg to 70deg; the should be at least 10ft square by 5tf

deep, and into this the spoil is washed by the giants, although it is better practice to feed the

material directly into the intake or ground section of the elevator, possible.

To prevent the throat of the elevator being choked and clogged up with large stones, it

is advisable to fit a grating or grizzly over the end of the sluice, which delivers the material to

the ground section or intake of the elevator, and the bars of the grizzly should be so spaced

that the distance between them 1 in less than the diameter of the throat. Liners reinforce the

entrance section and the throat, as these take the most severe wear.

4.7.1.1 Water required

It is usually stated that the elevator nozzle water will lift 0.5 times its own volume of

outside water, and thus half the supply of water under pressure is allotted to the elevator and

the remaining that the giants can excavate, and thus if 50% of the water supply is allotted to

the elevator and 50% to the giants, to obtain the greatest efficiency it is necessary to see that


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the parts of the elevator are correctly proportioned with regard to existing conditions, that is

,the sizes of the nozzles of the elevator and the giants should be such that they discharge these

amounts, in which case the elevator will lift all that the giants discharge.

4.7.1.2 Maximum height of lift

An approximate rule is that the natural head of water required for an elevator should

be about five times greater than the vertical lift required, which means that the maximum

height of lift would be about 20 of the effective head; in practice, on the other hand, it is

found that the maximum height of lift varies from about 15 to 17% of the effective head at the

nozzle of the elevator.

However, no definite rule can be given for the ratio between the effective head and

vertical lift, as the latter is dependent to some extent on the size of gravel, the slope of the

upraise pipe and the pressure, as a high pressure or head of water is much more effective than

a low head.

4.7.1.3 Capacity of elevators

There again no definite rule can be given for determining the elevator capacity, that is,

the quantity of material and water lifted, as this aloe varies according to the ratio between

effective head and vertical lift, the volume and head of water available, the duty of the giant

water the latter being the most vital factor in calculating the size of the elevator.

At the very outside the solid material does not form more than 5% of the entire weight

(water and gravel combined), lifted by the elevator, and in practice it is generally reckoned at

from 2% to 3% and so it is usual, when calculations the size of an elevator, to consider only

the water capacity.

4.7.1.4 Efficiency of elevators

The efficiency of hydraulic gravel elevators is very low, and varies from 10% to 20%.

PEELE (Mining Engineers' Handbook by ROBERT PEELE) gives the following

formula for calculating the efficiency of elevators:

E = H (62.4 W + S) / 62.4 N (H1 H)


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where, E = % efficiency,

H = Height of lift, in feet,

H1= Effective head, at nozzle of elevator,

W = Cubic feet per minute of water (giants, seepage, etc.)

S = Weight of gravel elevated per minute, placer gravel being taken as

weighing 3000lb per cuyd.

N = Water discharged through the elevator nozzle, in cubic feet per minute.

4.7.2 Sluice boxes for hydraulic elevators

The width of sluice boxes to be employed in conjunction with hydraulic elevators has

already been in, and these should be adhered to as closely as possible. The line of sluice boxes

is usually short, as the gravel has already been thoroughly disintegrated during its passage up

the upraise pipe of elevator.

Long ridge (Hydraulic Mining by C.C LONGRIDGE) gives the following grades for

elevator sluices:

1stBox of 12ft 2 in

2nd

Box of 12ft 3 in

3rdBox of 12ft 4 in

4th

Box of 12ft 5 in

5thBox of 12ft 6 in

The bedrock ditches or sluices, which feed the excavated material to the elevator sump

or ground section, should have a grade of from 8% to 15%, a 10% to 12% grade being

common.


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CHAPTER 5

METHODS FOR UNDERGROUND MINING OF COAL DEPOSITS

5.1.Basic Factors Influencing the Choice of Methods for Mining Coal Deposits

The principal factors that influence the choice of methods for coal occurrences are

mentioned below. A number of features mentioned below are of importance in working ore

deposits as well.

(1) Shape and Thickness of a Deposit

In the ideal case, a mineral bed has a regular thickness. In coal mining practice,

however one frequently encounters considerable variations in thickness caused by

disturbances of genetic and tectonic nature. In selecting mining methods due account shouldbe taken of frequent changes in thickness, roughness of the bottom and back, swells and

petering out of beds.

In mining thin and medium-thick beds full-seam extraction is ordinarily practiced,

while in working thick beds either slicing or full-seam extraction is preferred. In thick beds

the mined-out space often has to be stowed with mine-fill.

The minimum working thickness of a seam fluctuates within a wide range and

depends upon the angle of dip, characteristics of the bed, machinery used and the nature of the

reserves in a given basin. The maximum working thickness of coal seams in different

coalfield varies from 3 to 30 metres and more.

(2) Dip Angle

An important factor in choosing mining methods is the angle of dip. With a heavy

pitch, for example, the mineral moves by gravity and this is always taken into account in

selecting the system of mining. In coalmines, level, gently sloping (up to 25) inclined (25 -

45), and steeply pitching (45 - 90) seams are distinguished.

Frequent changes in the dip angle within one and the same deposit tend to complicate

its mining.


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(3) Barren Intercalations

Sometimes mineral beds contain barren intercalations. Their presence complicates

mining operations, since the barren rock has to be thrown into the worked-out space, or used

as mine-fill to avoid an increase in the ash content of the coal. To separate intercalations less

than 5 cm in thickness is practically impossible and the gangue contained therein becomes

mixed with coal. Considerable complications are caused by intercalations of barren rock

susceptible to self-ignition. Rock from such intercalations should never be left in the mined-

out space.

(4) Properties of Coal

The basic properties of coal include rigidity, schistosity and cleavage, susceptibility to

self-ignition, chemical composition and technological properties, size characteristics after

breaking, firedamp evolution.

5.1.1 Rigidity of coal

By rigidity of coal is meant its resistance to mechanical effects. The rigidity of coal is

dependent upon its hardness, toughness, mineralogical composition and structure, presence of

cracks and cleavage.

The rigidity of coal greatly influences the selection of undercutting and breaking

methods.

5.1.2 Schistosity and Cleavage

Sedimentary rocks can be split along parallel planes of bedding. This property is

termed schistosity. In some rocks schistosity is related to the time of rock formation (primary

schistosity). Rocks often exhibit schistisity of secondary origin.

A system of joints of secondary origin, formed as a result of tectonic processes, which

have transformed solid monolithic rock into schistose rock, is called cleavage. As distinct

from primary schistosity, cleavage is generally non-parallel to the bed surface. The spatial

position of cleavage planes is usually characterized by the angle between the cleavage

orientation and the strike of the bed.


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The cleavage planes and their spatial orientation strongly influence the choice of

mining methods. Cleavage in the roof and its orientation in space are of great importance for

roof stability. In the working of steeply dipping, and particularly of thick beds big lumps of

coal may become detached along the cleavage surface. Changing the sequence of coal

extraction and using the appropriate support at the face can avoid this hazard.

5.1.3 Susceptibility of coal to self-ignition

This property of coal strongly affects the choice of mining methods. Seams of self-

ignition coal should be extracted with the least possible losses and with close packing of the

mined-out space, since with a rise in temperature the abandoned coal rapidly oxidizes and this

often results in its spontaneous combustion. Fire hazards increase in mining thick beds.

Extraction of coal contains minimum amounts of volatile matter (anthracite) present less the

hazard.

5.2.Chemical composition and technological characteristics

The chemical composition and technological characteristics of coal are established on

the basis of elementary and technical analyses. The former serves to determine the content of

carbon, hydrogen, oxygen and nitrogen. The latter is of paramount commercial importance,

for it indicates the content of moisture, ash, sulphyr, phosphorus, the yield of volatile matter

and the calorific value of the coal (this is largely dependent upon the content of combustible

elements-carbon and hydrogen). In mining coal measures the grades of coal are taken into

consideration when choosing the order of extraction and the methods for mining individual

seams. The self-igniting properties of coal, which largely depend upon the content of sulphur,

are taken into account in selecting mining systems and fire-fighting measures to be applied in

the course of stoping.

5.2.1 Size of coal

The size and number of lumps in the extracted coal are of great importance for certain

grades of coal (anthracite, gas coal). This should be reckoned with in selecting the methods of

coal breaking and transportation.


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5.2.3 Gas-bearing capacity of a deposit

Mining of coal beds may be attended by the evolution of various gases, such as

methane (fire-damp), carbon dioxide, carbon monoxide, hydrogen sulphide, hydrogen and

gases released by blasts. Methane is released most frequently and in appreciable amounts.

Methane evolution from coal and wall rocks takes various forms, such as placid outflow,

release with audible sound effects (fumaroles), and sudden noisy out rushes with outbursts of

fine coal. Sudden out rushes are the most dangerous, since vast amounts of gas and coal (from

several tons to several hundreds and even thousands of tons) are ejected.

Sudden out rushes begin at a mining depth of 200 - 300m. They occur in many

coalfields and regions.

The discharge of methane is one of the prime factors to be considered in choosing

underground mining methods. Being lighter than air, methane tends to fill the space of

ascending mine workings. One of the major means to reduce the concentration of methane is

the application of mining methods involving a minimum of dead faces and upraises.

It should be borne in mind that in mining coal deposits there may be sudden releases

not only of methane, but also of carbon dioxide.

5.2.4 Properties of Wall Rocks

In the conditions of mining coal occurrences the wall rocks are commonly made up of

clayey, sandy and coaly shales, sand-and limestone. Knowledge of wall rock properties is

needed to enable one to solve problems linked with the excavation and timbering of mine

workings, rock pressure control and the application of the appropriate mining system.

5.2.5 Strength of rocks

By the strength of rocks is meant the ability of a solid body to withstand the action of

definite forces producing a particular deformation. It is characterized by ultimate strength of

the rocks (in kg/cm2) in tension, compression, bending and shearing. These are determined by

laboratory tests of rock specimens. The strength characteristics of rocks depend upon the size

of the grains making up the rock, the binding agent, porosity, water-content, and also the size,

shape and stress condition of the specimen. Coarse-grained and porous rocks are softer than


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fine-grained ones. Much attention should be paid to establishing the strength of rocks, for this

determines their stability, which is of prime importance in mining.

5.2.6 Stability of rock exposures

The deformation of a rock exposure may for a long time remain within the elastic

limits. In such instances the exposure is regarded as stable and no timbering is normally

needed. Should the deformation of a rock exposure surpass the elasticity limit the exposed

rock surface changes its position. In this case the state of the exposure is designated as

unstable. Deformation of such exposures often ends in sliding or caving of individual sections

and pieces of rock. Exposures of friable rocks liable to collapse without any discernable

deformation are also unstable.

In assessing the stability of rocks one should consider their strength and, in particular,

the presence of jointing, both of primary and secondary origin.

The stability of wall rock exposures in the roof and floor is one of the major factors to

be considered in selecting the system of mining, rock pressure control and design of support

to be used in the stoping.

In drawing up full characteristics of wall rock properties account must be taken of the

joint effect of such factors as the depth of occurrence, the presence or absence of water, the

nature of the rock sequence, the thermal conditions and tectonics disturbances with and

without discontinuity of the rock generally lower its strength.

5.2.7 Mutual Arrangement of Seams in a Series

In mining closely spaced beds a definite sequence of extraction and a more-complex-

than-usual support of the mined-out space should be adopted. These two factors require the

application of strictly specified mining methods.

5.2.8 Water-bearing Capacity of a Deposit

When exploring a deposit its water-bearing capacity should be ascertained. In the case

of water abundance means should be provided for the drainage of individual sections,

protective pillars left near water bodies which cannot be drained for some reason, a safe

method of roof control adopted, etc.


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5.3.Classification of Methods of Mining Coal Deposits

The various methods of mining involve production faces that are either long (long

wall) or short (short wall, or room and pillar). The method of mining controls the methods of

development of the take and the districts, and the mechanization of the production faces.

Until mining machines were developed, when handwork predominated, short faces

were preferred.

About the end of the last century and during the early 1900s, work on underground

mechanization began, first came the mechanization of coal winning, by cutting, breaking and

delivering the coal along the face. The first coal cutting machines and conveyors appeared. So

as to utilize these machines more fully, European countries began to turn towards long faces

and away from short ones.

The subsequent development of mining machines favored mining by long faces even

more. The transition to the long face became established as the aim to reduce the volume of

development work, since mechanization in developments lagged behind that in production

faces.

The advantage of short wall faces is the lightening of the face supports and the

absence of special caving supports as a consequence of roof control by mineral pillars with

subsequent caving or lowering of the roof after the production face has advanced some

distance.

With appropriate equipment and organization, the elimination of the laborious takes of

roof control makes for a high output per man shift or short faces.

Some disadvantages of short wall work are the considerable loss of mineral, which

may reach and exceed 40 per cent, and the difficulty of ventilation.

Notwithstanding the considerable number of mining methods in existence, they can be

reduced to the following main types.

(1) Methods of working with long faces:

(a) Pillar methods, either

(i) Long pillars (long wall retreating), or

(ii) short pillars.


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(b) Long wall advancing methods, either

(i) With horizons divided into sub-levels, or

(ii) Without subdivision of the horizon into sub-levels (horizon-face)

(c) Composite methods of working

(2) Methods of working with short faces:

(a) Room methods

(b) Room and pillar methods.

In the classification of methods of working, the type and volume of the development

work needed in any method can be included.

5.3.1 Methods of Working with Long Faces

In mining by long wall retreating or long pillars the developments cutting the districts

into long pillars are driven ahead of the production faces. The work on the upper horizons

generally precedes that on the lower horizons.

In long wall advancing roads do not have to be driven ahead of the face, except that

the haulage road, to provide a safe reserve and convenient loading of the coal from the face,

should be not less than 50 m ahead of the coalface.

Ripping and building a road through the gob behind the coalface usually form the

brake inclines. The lower sub-levels (faces) are worked before the upper sub-levels.

The paired-panel method is a combination of the long pillar method with long wall

advancing. The odd-numbered sub-levels worked by long wall advancing, the even-numbered

sub-levels by long wall retreating.

The long pillar (retreating) method of working can be used with any angle of dip and

any thickness of seam when the whole thickness is taken, or when thick seams are extracted

by sloping slices, discussed later.

Long wall advancing can be used for taking seams of the worse categories of

gassiness, also for seams thinner than 1.3 m with roof and floor of average strength or better.


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5.3.1.1 PILLARS METHODS

Pillar methods of mining are employed chiefly in seams of medium thickness with

low-rigidity wall rocks. The working of such seams is attended by intensive roof caving.

Under such conditions it is technically difficult to maintain workings in the mined-out area,

and for this reason in pillar in an intact solid mass of coal, the undermined sections of the

workings being gob up.

(i) Long Pillars (long wall Retreating)

Relative to the seam, the gate roads which form the pillars can be driven to the strike, to

the rise or diagonally (rarely, with gentle angles of dip, to the dip). Depending on the direction

in which they are driven they may be:

(1) Long pillars (or panels) to the strike,

(2) Long pillars (or panels) to the rise,

(3) Diagonal long pillars (or panels)

Independently of the method of forming the pillars or panels, the mineral can be won

by working in any direction. For example, long pillars, or panels, along the strike can be

worked by faces laid out either to the dip, or to the strike, or diagonally.

(1) Long Pillars (or Panels) to the Strike

(a) In Gently Sloping Thin and Medium-Thick Seams

As pointed out above, the faces in the different horizons, taken by long pillars, or

panels, can move towards the shaft (retreating) or away from the shaft (advancing). In the

retreating method before the production faces start, the lower main lateral (the haulage level)

and the return airway must be driven for the full length of the flank. In the advancing method,

the haulage roads have only to be driven for the length of the district.

Each district is served by its brake incline, driven before the faces start producing

material.

The district brake inclines can be single-sided or double-sided.

An advantage of double-sided inclines is that it is possible to double their output of

mineral compared with single sided inclines, particularly if conveyors are used within the area

served by the incline.


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Double-sided inclines are rarely used in extraction by horizons; each individual

instance needs to be based on technical and cost advantages.

Long panels to the strike begin to produce after all the development roads have been

driven to the end of the district that is after the panels have been cut and the headings to open

the faces have been driven.

Extraction of the separate panels (sub-levels) should proceed regularly with the upper

sub-levels 10-15 m ahead of the lower ones. Unworked pillars surround the district brake

inclines because the first sub-level taken is the upper one, which reduces their maintenance.

Within each district the faces can advance or retreat.

Advancing faces move in the same direction as the development work, and the mineral

is delivered by the middle gate to the forward brake incline.

When the faces retreat towards the brake incline, their direction of movement is

opposite to that of the developments and the mineral is delivered from the faces along the

middle gate to the rear brake incline.

In the rear brake incline there are two brake haulages, one of them delivering coal tot

he main lateral from the faces the other delivering it from the developments. In the forward

brake incline, there is only one brake haulage.

Using the rear brake incline, the losses in pillars around the brake incline are usually

larger than when using the forward brake incline.

The great advantage of using the rear brake incline with a general advancing direction

is the smaller distance of transport compared with the forward brake incline. With the latter,

the entire mineral is hauled along the gate road away from the shaft, down to the main lateral

where the haulage direction is reversed.

With a general retreating direction, the work of transport to the forward brake incline

will be smaller than to the rear brake incline. Another advantage of using the forward brake

incline is that there is no need to drive headings to open each face because after the extraction

of the panels between the brake incline and the man way, the faces move ahead directly to the

new district. If full extraction of the panels is impossible work in the subsequent district can

begin directly from the man way.

Roof control is also easier when using the forward brake incline because the greatest

danger of falls in the face occurs at the moment of the first caving of the main roof.. This


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takes place after the face has moved some distance away from the point where it started.

Using the rear brake incline, these conditions would occur in every district when the face is

opened up, but using the forward brake incline they occur once only when the first face of the

horizon is opened.

Double-sided brake inclines are driven not at the boundary of the district but at its

center, the gate roads being driven from the brake incline in both directions to the boundary of

the area served by it.

The faces for each new double-sided brake incline start by the driving of a heading at

the limit of the area worked by this incline, and retreat towards the incline. By this means half

the mineral reserves are delivered along the middle gate to the rear break incline, and half to

the forward incline.

The increased operating expenses of this method are somewhat compensate by the

reduction in the number of brake inclines per horizon through the increased length along the

strike of the area worked by the brake incline.

In extracting gently sloping seams by long panels along the strike, the slope length of

the horizon is from 250 to 450 m or more.

In choosing the height of the sub-level (the length of the face) many factors must be

considered, the main ones being: the cost of driving the middle gates and their maintenance,

the type of mechanization for getting and transporting the mineral to the middle gate, the ease

with which men can travel to the face, and the delivery of supplies.

From the viewpoint of the cost of driving and maintaining the middle gate roads, it is

desirable to increase the height of the sub-level because this reduces the number of middle

gates.

So as to use coal cutters and cutter-loaders rationally and to improve the organization

of the work, the height of the sub-level is generally chosen to equal the machine's capacity in

one or two shifts.

From the viewpoint of delivery of the mineral from the face to the middle gate, the

face length should logically be equal to the conveyor length.

Roof control does not limit the height of the sub-level. In other words, for any face

length a rational method of roof control can be found.


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Separating the air currents for the air currents for the production faces and

development faces usually ventilates long panels.

(b) In Inclined or Steep Thick Seams

Long pillars to the strike are used for working seams up to 4.5 m thick, if they are

inclined or steep.

Depending on the seam thickness and the angle of dip, the horizon is divided into two,

three, or sometimes four sub-horizons with faces 25-30 m long depending on the strength of

the coal together with a parallel gate road joined to it periodically by crosscuts in the return

airway by airways every 50-70 m along the strike. In gassy mines this spacing has to be

reduced to 30-40 m.

In the former haulage roads of the upper horizons cannot be used as return airways,

below them, 4-5 m down the dip, special airway are driven, joined by inclines and crosscut to

the upper return airway.

The face is timbered with rows of bars down the tip, using half-rounds or boards 3-4 m

long. The bars are butted to their neighbors and held by props at their ends.

Depending on the type of roof and floor, the spacing between props in a row is 1-1.2

m, with 1-1.3 m between rows. The props are set in sinking, in the floor; if the floor is weak;

sills are placed in the sinking, with notches in them for the props.

Full caving using breaker props mainly performs roof control. The breaker props are

usually left in but the face timbering is partly withdrawn.

(2) Long Pillars (or Panels) to the Rise

In long pillar mining to the rise development work is similar to that done in long pillar

working on strike, except that each sublevel is divided into pillars to the rise by special raises.

Recovery of each pillar is started from the top by driving butt entries. From production faces

coal is conveyed via the rise headings (raises).

This, and the fact that short walls draw coal, explains why it is possible to use this

system when wall rocks are poor. The pillars are blocked out gradually, as the need arises.

This method has many disadvantages: (1) it requires many openings to be driven with

narrow faces; (2) production places are father short this making the use of mine machines


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difficult; (3) ventilation is an extremely complex affair; (4) the system is inadmissible in

gassy mines in view of the danger of methane accumulating in rise headings; (5) scattered

working places; (6) high coal losses.

(3) Diagonal Long Pillars (or Panels)

The method of working by long diagonal pillars or panels is used for working nearly

level seams (2-3).

In this case it is possible to find the anglebetween the strike line and the direction of

the diagonal of gate road, along which its inclination will reduce to 0.5-1(a slope of 0.9 to

1.7 percent). Under these conditions there is no need for a brake.

It is easy to establish the relation between the slope angle of the diagonal gate road,

the angle of dip of the seam and the angle between the diagonal gate road and the strike.

Sin = sin sin

Because of the smallness of and for convenience, we can write this equation

Tan = sin sin

The angle of the diagonal pillars should not be too acute because otherwise the coal at

the points of the pillars will crush, and the roadway supports will therefore suffer extra load at

this road junction. If it is reduced below 50-60, the sharp point of the coal pillar should be

taken off for 3-4 m.

Strike

Dip

Fig. Diagonal long pillar

Coal seam


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(Diagram of the relation between the gradient of a diagonal level, the dip angle of the

seam, and the horizontal angle between the strike line and the level.)

(ii) Short Pillars

If the mining conditions are such that faces first opened on the long panels cannot be

taken out to the full length along the dip, each long panel can be split with level roads and

headings into, smaller parts, or short panels.

Work begins in exactly the same way as with the system of long panels. After the

driving of the gates and brake inclines each sublevel is cut into short pillars by headings and

slits without ripping or dinting. The dimensions of these pillars are from 8 x 8 m to 20 x 20 m,

depending on the strength of the ground.

The large amount of development work, and narrow work generally, makes the use of

short pillars irrational. It is not practicable in gassy mines because of the large number of

dead-end workings in the faces of the headings and slits, which methane can accumulate. In

addition, it is exceedingly difficult to achieve integrated mechanization of the work in

working short pillars.

Short pillar mining can be used also in a different way. With well laid-out ventilation

in the slits cutting the panels into short pillars, it is possible to work without headings. When

this occurs, the short pillars are worked by small buttocks each side of each slit.

The advantage of this way of working, by comparison with short pillars proper, is the

smaller quantity of secondary development work because of the reduced number of headings.

(b) LONGWALL ADVANCING METHODS

(I) With Horizons divided into sub-levels

(In Gently Sloping Seams)

The long wall advancing method of working with the horizon divided into two sub-

levels, and the face of the lower sub-level some 20-30 m ahead of the upper one. The best

ventilation is provided under these conditions, and if there is any water it flows down the gob

and not into the face.


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The slope height of the sub-level (the face length) is chosen from the same

considerations as with the method of long panels.

In thin and extremely thin seams, the face length is usually 100- 150 m. occasionally it

can be shorter or longer because of local geology or organization of work.

When the horizon is divided into sub-levels, coursing ventilation of all the sub-levels

in sequence by one air current from the haulage road is forbidden in gassy or dusty mines, as

for the method of long panels.

(ii) Without Subdivision of the Horizon into Sub-levels (Horizon face).

(In Gently Sloping Seams)

If the seam is more than 0.8 m thick it is possible to work the coal without division of

the horizon into sub-levels. A single face can work the horizon at least 150 m long and the

method is described as the horizon-face.

With this method of working, it is characteristics that there is no middle gate between

the bottom (haulage) gate and the return airway, nor is there any brake incline.

With the method of working by a horizon-face, the ventilation is exceedingly simple.

The air current arriving through the haulage road enters the face, passes through it and out by

the return airway. The face of diverges for the haulage road some 50 m ahead, usually causes

no difficulty. The safety rules, however, must be observed; the speed of the ventilating air in

the face should not exceed 4m/sec and the methane content of the return air should not exceed

1 per cent.

In these methods, the long wall face travels along the strike of the seam, and this is

typical of long wall workings.

If the coal has a cleat diagonal to the strike, the face may also be diagonal to make it

easier to work the coal. This method of working can be called long wall working to the strike

with a diagonal face.

Sometimes the diagonal layout of the faces is necessitated not by the need to work a

coal with a certain cleat, but by the need to angle the face to avoid the roof cracks which are

parallel to the face.


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(C) COMPOSITE METHODS OF WORKING

Methods of working seams include so-called composite methods of working, using

some features of two systems, the commonest being the method of paired panels.

In this method, the horizon is divided into panels or sublevels, the even-numbered

panels being worked by long wall advancing from the brake incline to the boundary of its

area. When the faces reach the boundary, opening headings, are driven into the odd-numbered

coal panels and these are worked retreating the long panels, towards the brake incline.

The gob in the even-numbered panels is packed with stone obtained from the ripping

of the gate roads, and the odd-numbered panels are caved. This method is used in very gently

sloping seams up to 0.7 m thick where packing of the stone from the ripping into the rise side

face packs is not too laborious.

During the working of the sub-level from the brake incline to the boundary of the

district, difficulties may occur which are characteristic of long wall advancing. Coal pillars

are on one side of the returns and intakes, and on their other sides between them is a gob, a

layout which is usually unfavorable for the maintenance of the roads. The duration of

extraction of the horizon using paired panels is long. Besides, extraction of one thin seam by

only one face per gate is unfavorable also for transport. These disadvantages limit the

application of the system of paired panels. But it has been fairly widely for working thin

seams.

An advantage of this method of working is that it is easy to ventilate the faces

separately without driving additional roads. Besides, it can be used for preliminary drainage

of the area.

(2) Methods of Working with Short Faces

(a) ROOM METHODS

For working horizontal or very gently sloping seams, the mine take is divided up into

panels bounded by pillars. The panels are taken by retreating; using a large number of

simultaneous short faces (rooms) giving a high total output.

The pillars between entries are from 10 to 15 m wide, the entries being driven 3-4.5 m

wide on a narrow face. In thin seams the entries may be ripped with considerable delay, up to


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100 m behind the faces. The ripping are either piled up along one side of the entry or are

dumped in an old room nearby.

In inclined seams, the rooms are worked in groups or in sequence from the panel

entries (cross-entries) or directly from the main entries. The pillars between rooms are not

extracted. For ventilation and the traveling of machines and men from room to room, the

pillars between rooms are slit with through connections.

The rooms are driven at 45to the cross entries and the through connections are at 45

to the rooms. The width of the rooms and pillars is from 6 to 12 m, a room being 100-120 m

long. All the winning and loading work is completely mechanized with coal cutting and

loading work is completely mechanized with coal cutting and loading machines or cutter-

loaders mounted either on crawler track or rubber tires.

The rooms are supported with roof bolts.

The coal is conveyed from the rooms by scraper conveyors or shuttle cars.

(b) ROOM AND PILLAR METHODS

The room and pillar method differs from that of because the pillars between the rooms

are partly extracted. Under particularly good conditions they are completely extracted by

retreating after all the rooms have been taken.

The losses of coal in this method are smaller than in the room method but coalwinning and the roof support are more complicated; therefore the output from the men is

lower.

In the first variant, the rooms are located on both sides of the cross entries. In the

second, extraction of the second panel begins only after the face of the first panel has

advanced 30-40 m. With single-sided work, pillars must be left between the areas extracted,

increasing the losses.

The width of the rooms depends on the strength of the roof. In less strong ground the

rooms are as wide as the pillars between them, or a little narrower. In stronger ground the

pillars are narrower than the rooms.

The work in the rooms is usually mechanized in the same way as for entries driven in

thick seams.


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CHAPTER 6

METHODS FOR UNDERGROUND MINING OF ORE DEPOSITS

6.1.Fundamentals of Mining Ore Deposits

6.1.1 Distinctive Features of Working Ore Deposits

Most of the ore deposits represent lenses, stocks, domes, pockets, lodes or a

combination of ore bodies. The mode of occurrence is extremely irregular, the angle of dip

and the line of strike being liable to frequent variations. It is especially difficult to mine lode

deposits representing a set of thin veins with irregular spatial position and outline. In working

such deposits extensive prospecting is needed and its result should be brought in line with the

exploitation requirements. A specific feature of one deposits is the diversity of the properties

of the ore and host rocks and the complex composition of the ore. Both the ores and enclosing

country rocks vary from friable to very strong, from soft to extremely stable. The complexity

of composition is typical primarily of non-ferrous ore deposits that are usually polymetallic.

These ores often belong to distinct grades and classes, which need different methods of

dressing and processing (rich and lean, sulphide and oxidized, copper and zinc-copper,

chalcopyrite and pyrites, etc.).

By reason of high technological standards and the necessity of delivering ores of

different grades and classes, selective extraction in mining ore deposits is often a matter of

prime importance. In working very thin ore bodies, separate recovery of the ore and barren

host rocks, picking and sorting of gangue in underground workings and on the ground surface

acquire great importance.

6.1.2 Basic Mining Characteristics

In establishing factors influencing the choice of a mining method and in deciding other

issues pertinent to the working of ore deposits one should be guided by the accepted practical

characteristics of ore bodies, the ore and enclosing country rocks.

(a)Classification of Ore Bodies According to the Dip Angle

As regards the angle of dip, ore bodies are classified as: -


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- Horizontal and conventionally horizontal with a dip angle of 0-3(the maximum angle of 3

is adopted in conformity with the maximum permissible track gradient for electric locomotive

haulage of the muck directly from the stope area);

- Gently sloping ore bodies with a dip angle of 3-30(the ore in the stope area is hauled by

mechanical devices scrapers, conveyors);

- inclined ore bodies with dip angles of 30-45 (the ore in the stope area can be hauled by

simple arrangements, such as metal troughs and metal-lined chutes);

- steep ore bodies with dip angles of 45-90 (the broken ore can move in the stope area by

gravity.)

(b)Classification of Ore Bodies by Thickness

As concerns their thickness, the ore bodies are classed as -

- Very thin (up to 0.7-0.8 m thick);

- Thin (0.7-0.8 to 2 m);

- Medium thick (2 to 5 m);

- Thick (5 to 15-20 m);

- Very thick (over 15-20 m).

The classification of ore bodies according to their thickness is based on yhe most

frequently employed methods of excavating development openings, or on the direction of

stoping. Thus, in very thin ore bodies stoping and excavation of major development workings

(drifts, raises) are done with additional ripping of wall rocks; in thin ore bodies stoping is

carried on without ripping, while major development workings are driven with ripping of wall

rocks; in ore bodies of medium thickness the stoping and excavation of major development

openings are effected without ripping of wall rocks in mining thick ore bodies stoping is

usually done along the strike of the ore body; very thick ore bodies are worked across the

strike.

(c)Strength of Ores and Host Rocks

The ores and rocks are divided into five groups: loose, soft, friable, strong, and very

strong.


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For more precise assessment of the hardness of the ores and host rocks, the hardness

scale is used. The rock hardness figures established according to this scale may, whenever

necessary, be verified by compression tests of rock specimens of definite shape and size. The

International Bureau of Soil Mechanics in Prague recommended in 1961 that test specimens

be obtained by drilling out a core with a height and diameter of 40-45 mm, the parallelism and

flatness of its faces being strictly observed. The Bureau established a scale that includes 12

classes of rock strength from 30 to 3,000 kg/cm2.

In making calculations for drilling and blasting work and in setting up the relevant

standard characteristics one should use special classifications of rocks in respect to drillability

and blastability. More detailed information on the classification of rocks as to their hardness,

drillability and blastability may be found in textbooks on drilling and blasting.

In solving problems of rock pressure, drilling of the ground and other problems

relating to the mining of mineral deposits, great importance is now attached to the

classification of the physical properties of rocks.

(d)Stability of Ore and Enclosing Country Rocks

By stability is meant the ability of rocks not to collapse when exposed from below and on

the sides. The stability of rocks greatly influences the choice of a mining system. A

classification of the stability of ores and enclosing country rocks is presented below.

(i) Soft and very unstable ground not allowing even the slightest exposures in the roof and on

the walls of mine-openings.

Advance timbering is often necessary when marketing openings in such ground. Soft

grounds include loose, running and friable aqueous grounds. In selecting mining systems and

the sequence of working individual portions of a deposit it is very important to know whether

such ground is present in the hanging or foot walls of the occurrence, since special

precautionary measures should then be envisaged.

(ii) Unstable ground permitting very limited exposures of the roof and walls in the opening

immediately adjoining the working face.

In such grounds the back and walls of the opening should be supported right after its

excavation.

(iii) Medium-stability rocks allowing limited exposures of the roof and walls in the opening.


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The need for artificial support of exposed ground does not arise at once, but with the

passage of time and not always over the entire exposure area.

(iv) Stable rocks permitting extensive exposures of the roof and particularly of the walls, and

requiring support only at individual spots or at certain intervals.

(v) Very stable rocks allowing very large exposures of the roof and walls without any

artificial support.

Although this classification lacks a quantities index, it allows a practical approach to

the choice of a mining system taking into account the stability of the ground.

6.1.3 Methods of Stoping in Relation to the outlines of the Working Faces

In the practice of mining ore deposits diverse shapes of the stoping front, are adopted,

depending upon the particular mining and geological conditions and requirements for

efficient, convenient and safe breaking of the mined bulk. These following methods of stoping

are distinguished depending on the outline of the working faces; overhead-stoping,

underhand-stoping, and breast-stoping.

(a) Overhand Stoping

In overhand stoping the ore is broken by benches in ascending order, the lower

benches leading the upper ones. The miners work beneath the horizontal or sloping plane of

the bench. Horizontal, inclined or diagonal slicing may do overhand stoping. The adoption of

a particular version of overhand stoping is contingent upon the angle of dip of the ore body,

the mode of support used in the worked-out space, and the ore properties. The presence of

joints and their direction in the ore may prove essential. The latter factor should be taken into

account for safety reasons.

With an angle of dip exceeding 45-50overhand stoping allows the mineral to be moved by

gravity. Overhand stoping by horizontal slices is used most frequently.


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Bench

Slice

Fig. Overhand stoping

(b) Underhand Stoping

This is a mode of mining in which the stope is made in the shape of benches disposed

in descending order, with advanced stoping of the upper benches.

Slice

Fig. Underhand stoping

When drilling, the miner stands on the solid ore at the toe of the bench. Holes are

drilled downwards, except for reliever holes, which may be run horizontally or slightly

upwards. In underhand stoping getting the ore down requires some reshovelling.


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(c) Breast-Stoping

In this mode of extraction the stoping operation front runs along a single line. This line

may extend in any direction depending upon the geological conditions and the design

characteristics of the mining method.

Breast-stoping is commonly used in horizontal or slightly sloping ore bodies. The

conditions of application, the advantages and disadvantages of the basic modifications of

stoping are described when discussing the various mining systems.

Work-

out

Fig. Breast stoping

6.2.Classification of Methods for Mining Ore Deposits

Ore deposits are characterized by extremely great variety and complexity of their

working, therefore the number of methods and their variants used in the practice of mining

ore occurrences is quite considerable. The diversity of mining conditions and the great

number of existing systems complicate the elaboration of a simple and harmonious

classification of methods for mining ore deposits.

Most of the mining specialists believe that a classification of mining methods should

be based on the methods of supporting the stope area, or the condition of the latter. There is

no doubt that the mode of supporting the stope area and also its condition are the major

features characterizing a mining method, since they determine the general design of the

system and its technical and economic characteristics.


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A classification of methods for mining ore deposits based upon method of support is as

follows:

(1) Stopes naturally supported

(a) Open stoping

(i) Open stopes in small ore bodies.

(ii) Sublevel stoping.

(b) Open stopes with pillar supports.

(i) Casual pillars

(ii) Room (or stope) pillar (regular arrangement)

(2) Stopes artificially supported

(a) Shrinkage stoping

(i) With pillars

(ii) Without pillars

(iii) With subsequent waste filling.

(b) Cut-and-fill stoping.

(c) Stulled stopes in narrow veins.

(d) Square-set stoping

(3) Caved stopes

(a) Caving (ore broken by induced caving).

(i) Block caving.

(ii) Sublevel caving.

(b) Top slicing.

(4) Combinations of supported and caved stopes

6.2.1 Stopes naturally Supported

(a) Open stoping

Stopes naturally supported are those in which no regular artificial method of support,

although occasional props, cribs, or stulls may be used to hold local patches of insecure

ground. The walls and roof are self-supporting.


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Open stoping is applicable to deposits of strong, firm ore having strong, firm walls. In

the narrower deposits (10-15 m wide) the ore are often can be mined the full width in one

operation without the use of pillars. In wide deposits it usually becomes necessary to leave

solid pillars reduce the length of unsupported apan and thus prevent the failure of back or

walls.

The main condition for the use of the open-stope methods of mining is considerable

stability of the ore and the host rocks that would allow large areas of exposure to be left.

(i) Open stopes in small ore bodies

The simplest form of open stope is that in which the entire ore body is removed from

wall to wall without leaving any pillars. It is applicable to relatively small ore bodies, as there

is a limit to the length of unsupported span that will stand without support even in the firmest

and strongest rocks.

Plan Surface

`

Shaft

Ore boundary

Open stope

Bottom of ore

Fig. Open stoping without pillars

Bench

HeadinOpenstope

(ii) Sublevel stoping

The basic conditions for the use of the method are: considerable stability of the ore

and host rocks; dip angle of the ore body not less than 50; thickness of the ore body from 5-6

to 20-25 m (when it exceeds 20-25 m the rooms are extracted across the strike).


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The ore body is worked by horizons with an interval height of 50 to 75-100 m. With

strong ore and vary stable enclosing rocks it is possible and expedient to increase the level

interval to 70 m and more.

3

10

6 9

8

2

7

4 5

1

3

2 4

Fig. Typical variant of the sublevel method with a grizzly level

1- haulage drift; 2- raise; 3- grizzly drift; 4- ore chute; 5- dozing chamber; 6- sub-drift

7- finger raise; 8- open sublevel cross drift; 9- cut out raise; 10- ventilating drift

The horizon is broken up into blocks 50-60 m long each. With a high level interval the

length of a block often reaches 100-120 m. The length of the block should be in line, which

permissible exposure areas in the roof and walls of the room. With wide rooms the length of

the block is smaller. Each block is divided into a room and a pillar. The width of the pillar is

at least 6 m, the thickness of the floor pillar being 0.4-0.6 of the room width. The overall


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thickness of the floor (from the haulage elevation mark down to that of undercutting) varies

from 10 to 15 m with a grizzly level and 6 to 10 m with a scraper level.

Stoping in the room may be done in one direction or in two directions- from the centre

towards the boundaries. In the latter case the shortcomings due to the large length of the room

9for instance difficulties in delivering supplies to the stope) are eliminated and its yield is

increased.

Advantages and shortcomings

The basic merits of this system include: safety of operations, since the workers always

stay in mine openings (not in the stoped rooms); high labor productivity of face men thanks to

the possibility of using effective methods of breaking and drawing the ore; high intensity of

stoping operations; the possibility of uniform delivering of the ore to the haulage level.

The chief shortcomings of the method are: difficult and slow recovery of pillars; a

large volume of solid work; high losses and dilution of ore in robbing the pillars;

impossibility of sorting the ore.

The technical and economic characteristics of the systems of ordinary sublevel stoping

are as follows:-

Coef. of recovery - 0.95-0.97

Dilution factor - Depends upon gangue inclusions (in practice often from 0.05 to

0.10)

Output per face workers/ shift - 30-40 tons

(b) Open stopes with pillar supports

In open stopes with pillar support, leaving pillars of ore reduces the length of unsupported

span. These pillars may be of the irregular or casual type, their position and size being

determined by localized ground conditions or may be regular in size and arrangement,

conforming to a predetermined pattern.

(i) Casual pillars

This method is employed in mining gently sloping ore bodies 1-2 to 4-6 m thick with

stable host rocks. In an ore body of limited thickness the ore may have low stability.


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With the system breast-stoping in an ore body is done either along the strike or up the

raise, permanent pillars of lean ore or barren rock inclusion being abandoned in the course of

stoping. The pillars are sited at random and are of different size. In the absence of poor ore or

gangue inclusions in the solid ore pillars are sited in a regular fashion. In this case the room-

and-pillar method of mining is used.

In mining very valuable ores and when the deposit is less than 2 m thick artificial

pillars made of concrete and supplementary support of the roof (cribs or packing) may be set

up in place of natural ore pillars. Stowing is ordinarily employed in taking ore from thin

producing areas with ripping of the back and driving of advance headings. The variants of

breast-stoping with abandonment of permanent support pillars are distinguished by the mode

of ore delivery or the shape of the stope face, i.e. scraper or conveyor haulage, delivery on a

gravity incline, or delivery on rails to an ore-drawing chute made in country rocks, with

rubber-tired self-propelled mine-cars to a drift, with an underhand-shaped face.

Depending upon the angle of dip different modes of ore haulage are practiced; thus

rail transport and self-propelled mine-cars are used with a dip angle of up to 3, scrapper

haulage for 0-30and conveyor delivery for 0-15.

Stopes faced are shaped to conform to the thickness of the ore body, the technology

and organization of drilling and blasting operations. Benchless stoping is commonly

employed with ore bodies up to 3-4 m thick, underhand stoping with thicker ore bodies.

Advantages and shortcomings

The advantages of the system are: a small amount of development work; simple design

features; a considerable stoping area; abandonment of gangue or lean unsalable ore

inclusions; the possibility of sorting the ore directly in the stope area and retaining un-

inventoried ore reserves for future extraction.

The major shortcomings of the method are: high ore losses in permanent support

pi

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