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MI 31003 Underground Metal Mining Methods Lecture Notes

K.UMAMAHESHWAR RAO

Chapter 1

Salient features of Indian Mining Industry

1. The major contributors of mineral in the country are:

Table1. Share of key mining states on India’s mineral resources (Ministry of Mines, Government of

India; Ministry of Coal, Government of India, Indian Bureau of Mines, Centre for Monitoring Indian Economy -2006)

State Coal% Iron ore% Bauxite% Manganese

%

Lead-Zinc % Chromite

%

Jharkhand 29% 14% - - - -

Orissa 24 17 51 35 - 98

Chhattisgarh 16 10 - - - -

MP 18 - - 10 - -

AP (old) 7 7 21 - 1 -

Rajasthan - - - - 90 -

Karnataka - 41 - 29 - 1

Total 84 89 72 74 91 99

2. India produces about 87 minerals that include 4 fuel minerals, 3 atomic minerals, 10

metallic minerals, 47 non-metallic minerals and 23 minor minerals (including

building & other materials). India occupies a dominant position in the production of

many minerals across the globe.

3. There are close to 3000 mines in India. As per the records of 2010-11, of 2928 mines,

573 were fuel mines, 687 were mines for metals, and 1668 mines for extraction of

non-metallic minerals. Of the total number of about 90 minerals, the three key

minerals are coal, limestone and iron ore. There are 560 Coal mines (19% of total

number), 553 limestone mines (19% of total number) and 316 iron ore mines (11 % of

total number) bauxite (189), manganese (141), dolomite (116) and Steatite (113).

India ranks 3rd in coal production, 3rd in limestone production and 4th in iron

ore production, in the world as of 2010.

Table 2 .India’s Production Rank across Key Minerals – 2010 (Ministry of Mines, Government of

India; Ministry of Coal, Government of India, Indian Bureau of Mines, Centre for Monitoring Indian Economy -2006)

Mineral Application Total

Production

(‘000 tonnes)

India’s global

rank in

production

Coal Power, steel, cement 5,37,000 3rd

Limestone Cement, iron & steel, chemical 2,40,000 3rd

Iron ore Iron and steel 2,60,000 4th

Bauxite Transport vehicles, packaging, construction

materials

18,000 4th

Barite Oil and gas, paints, plastics 1,000 2nd

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Chromite Steel, dye & pigments, preservatives, refractory appli

cations

3,800 2nd

Zinc metal Iron & steel (galvanization), communication

equipment –as alloys)

750 4th

Managanese Iron & steel, packaging ( as alloy with

aluminium)

1,100 5th

Lead metal Paints 95 6th

Copper Electronics , architecture, alloys 161 10th

Aluminium Transport vehicles, packaging, construction 1,400 7th

4. Amongst the BRIC countries (Brazil, Russia, India and China), India is the least

developed in terms of per capita mineral consumption. As India’s per capita GDP

increases, its mineral consumption will grow at a rapid pace in line with the growth

witnessed in other emerging markets like China and Brazil.

5. Problems of sustainability of Indian mining industry:

Regulatory challenges:

There is no guarantee of obtaining mining lease even if a successful exploration

is done by a company. The mining licenses are typically awarded on a first

come first serve basis in principle but there is no transparent system.

Inadequacy of infrastructure: The inadequacy of infrastructure is related to

the absence of proper transportation and logistics facilities. Many of our mining

areas are in remote locations and cannot be properly developed unless the

supporting infrastructure is set up. For example, the railway connectivity in

most key mining states is poor and it has inadequate capacity for volumes to be

transported which adds to the overall supply chain cost. The government

foresees that steel production capacity in the country by the year 2025 will

increase to 300 million tonnes per annum. This would require Indian Railways

freight capacity to be around 1185 million tonnes, for only steel and its raw

material requirements.

Environmental clearance: A large percentage of mining proposals has failed to

get environmental / forest clearance from the Ministry of Environment and

Forests, Government of India.

Over and above these regulations, the mining companies also need to take the

local communities along, to ensure that they have the support of the ‘local’ side

for their projects. As a result, several projects are impacted with challenges by

way of opposition from local communities / NGOs, difficulties in land

acquisition, denial of clearances from the governing bodies, etc. A few instances

of some of the major projects that have been impacted in recent past are as

follows:

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a) Pohang Steel Company (POSCO’s) US$ 11 billion investment plan:

strong opposition from local people over land acquisition.

b) Vedanta’s proposed US$ 1.7 billion bauxite mining project in Odisha:

opposition by local community and eventual withdrawal of the forest

clearance

c) Utkal alumina project, which was a US$ 1 billion joint venture between

M/s. Hindalco (India) and Alcan (Canada) to mine and refine bauxite:

delayed by more than a decade due to challenges in land acquisition

d) Uranium Corporation of India Ltd., UCIL’s two mining projects worth

US$ 200 million and US$ 225 million in Meghalaya and Andhra

Pradesh respectively: opposition from local communities and

organizations on the grounds of likely effects of radiations on human

health and environment

6. Non-metallic mineral: The resource base of industrial / non-metallic minerals in

India is adequate except for Rock Phosphate, Magnesite and Ball Clay, for which the

estimates show decreasing reserves. In fact, country is deficient in fertilizer minerals

and heavily depends upon imports. Based on the industry these minerals find use in,

they are grouped under four categories

A. Fertilizer Minerals

1. Rock Phosphate 3. Sulphur and Pyrites

2. Potash

B. Flux and Construction Minerals

4. Asbestos 7. Gypsum

5. Dolomite 8. Wollastonite

6. Fluorspar 9. Non-cement grade limestone

C. Ceramics and Refractory Minerals

10. Quartz and other silica minerals 15. Pyrophyllite

11. Fireclay 16. Kyanite

12. China clay and Ball clay 17. Sillimanite

13. Magnesite 18. Vermiculite

14. Graphite 19. Non-metallurgical bauxite

D. Export Potential Minerals

20. Barytes 23. Mica

21. Bentonite 24. Talc, Soapstone and Steatite

22. Fuller’s Earth

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7. Mining of granite, marble, sandstone of building material quality (Chunar sandstone),

slate, barite, etc.; are classified under small scale mining sectors in the country

Chapter 1

1.0 Formation of ore deposits/ ore genesis

1.1 Introduction

The geological environment, the earth s has been subjected to various activities and as a

consequence it undergoes a cyclic change through a number of stages such as :

1. Erosion and planning (running down of mountains)

2. Weathering Stage, formation of sedimentary rocks

3. Sedimentary stage. burial in the deep crust –

4. Plutonic stage. When molten rock solidifies within pre-existing rock, it cools slowly,

forming plutonic rocks with larger crystals.(Plutonic – meaning deep underground; it

refers to the hydrothermal process where igneous rocks are formed by solidification at

considerable depths)

5. Orogenic stage –a stage characteristic of forces or events leading to large structural

deformations (folding, faulting, mountain building and igneous intrusions) of earth

lithosphere (crust & uppermost mantle) due to tectonic activity.

2. Concepts of Genesis of Ore

Ore genesis theories generally involve three components: source, transport or conduit, and

trap. The genesis of ore deposit is divided into internal (endogenic) and external (exogenesis)

or surface processes. More than one mechanism may be responsible for the formation of an

ore body.

Source is required because metal must come from somewhere, and be liberated by

some process

Transport is required first to move the metal bearing fluids or solid minerals into the

right position, and refers to the act of physically moving the metal, as well as

chemical or physical phenomenon which encourage movement

Trapping is required to concentrate the metal via some physical, chemical or

geological mechanism into a concentration which forms mineable ore.

The various theories of ore genesis explain how the various types of mineral deposits form

within the Earth's crust. Ore genesis theories are very dependent on the mineral

Syngenetic - A deposit formed at the same time as the rocks in which it occurs.

Ex. Banded Iron Formation

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Epigenetic- A deposit introduced into the host rocks at some time after they were deposited

Ex. Valley-type Deposits

GENESIS OF ORE DEPOSITS

Origin Due to Internal Processes

Magmatic

Segregation

Separation of ore minerals by fractional crystallization during

magmatic differentiation.

Settling out from magmas of sulfide, sulfide-oxide or oxide melts

which accumulate beneath the silicates or are injected into country

rocks or extruded on the surface.

Pegmatitic

Deposition

Crystallization as disseminated grains or segregations in

pegmatites.

Hydrothermal Deposition from hot aqueous solutions of various sources.

Lateral Secretion Diffusion of ore and gangue forming materials

from the country rocks into faults and other structures.

Metamorphic

Processes

Pyrometasomatic (skarn) deposits formed by replacement of wall

rocks adjacent to an intrusive.

Initial or further concentration of ore elements by metamorphic

processes.

Origin Due to Surface Processes

Mechanical

Accumulation

Concentration of heavy minerals into placer

Sedimentary

Precipitation

Precipitation of certain elements in sedimentary environments.

Residual Processes Leaching of soluble elements leaving concentrations of insoluble

elements.

Secondary or

Supergene

Enrichment

Leaching of certain elements from the upper part of a mineral

deposit and their reprecipitation at depth to produce higher

concentrations.

Volcanic Exhalative

Process

Exhalations of sulfide-rich magmas at the surface, usually under

marine conditions.

2.1 Spatial Distribution of Ore Deposits

It is considered that in certain periods of geological time scale, the deposition of a metal or

group of metals was pronounced; and also that specific regions of the world possess a notable

concentration of deposits of one or more metals.

Mineral deposits are not distributed uniformly through the Earth's crust. Rather, specific

classes of deposit tend to be concentrated in particular areas or regions called metallogenic

provinces.

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2.2 Mode of Formation

As hot (hydrothermal) fluids rise towards the surface (magma charged with water, various

acids, and metals in small amounts) through fractures, faults, brecciated rocks, porous layers

and other channels (i.e. like a plumbing system), they cool or react chemically with the

country rock.

Some form ore deposits if the fluids are directed through a structure where the

temperature, pressure and other chemical conditions are favourable for the

precipitation and deposition of ore minerals. The fluids also react with the rocks they are

passing through to produce an alteration zone with distinctive, new minerals.

2.2.1 Characteristic types of hydrothermal ore formations

Cavity Filling

The hydrothermal fluid fills in the cavities within the country rock and based on the shape of

solidified ore mineral several names have been attributed to the ore body shape, such as:

The cavity filling deposits are loosely termed as vein deposits Eg. gold, silver, copper and

lead-zinc. Veins range in thickness from a few centimeters to 4 meters. They can be several

hundreds of meters long and extend to depths in excess of 1,500 meters.

The process of cavity filling has given rise to a vast number of mineral deposits of diverse

forms and sizes. The Vein deposits resulting from cavity filling may be grouped as follows:

fissure veins, ( it is a tabular ore body that occupies one or more fissures: two

of its dimensions are much greater than the third)

shear zone deposits, ( thin sheet like connecting openings of a shear zone)

stock-works, (interlacing network of small ore bearing veinlets traversing a

mass of rock.

saddle reefs,

ladder veins, and

replacement veins or veinlet’s

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Fig .1 Various fissure veins: (A). Chambered vein; (B). Dilation veins; (C).Sheet veins;

(D). En-echelon vein (E). Linked vein

Fig.2 (a) Stockwork

Fig.2(b). Stockwork of a sulphide ore body

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Fig 3(a) Saddle reef

Fig.3(b). Bendigo Goldfield, Victoria, Australia

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Fig Ladder vein deposit.

Ladder veins are short, rather regularly spaced, roughly parallel fractures that traverse dikes

(tabular bodies of igneous rock). Their width is restricted to the width of the dike, but they

may extend great distances along it. Ladder veins are not as numerous or important as fissure

veins.

Questions:

Q1. What are the salient features of Indian Mineral industry?

Q2. Discuss the challenges of sustainability of Indian Mineral Sector?

Q3. Describe the geological processes involved in the formation of mineral resources.

Q4. Explain the characteristics and geometry of hydrothermal ore formations?

Q5. Geometric Measures of an Ore body

Axis of ore body: line that parallels the longest dimension of the ore body.

Pitch (Rake) of ore body: angle between the axis and the strike of the ore body

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ORE DEPOSITS and the Tectonic Cycle

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Lecture 2: Economic analysis for the assessment of viability of a

mineral resources sector

The first step of assessment whether a mineral deposit under consideration is viable under the

existing techno-economic conditions is to prepare a detailed feasibility report of the mining

project

Feasibility Report

A feasibility study is an evaluation of a mineral reserve to determine whether it can be mined

effectively and profitably or not. It includes the detailed study of reserve estimation, mining

methods evaluation, processing technique analysis, capital and operating cost determination

and the process effect on environment.

The feasibility study can be considered into two stages: prefeasibility studies and detailed

feasibility. Both stages are similar in term of content. The difference exist in the accuracy and

time required to perform the studies.

Detailed Feasibility Report:

This is the most detailed study to evaluate whether to proceed with the project. It is the basis

for capital estimation and provides budget figures for the project. It requires a significant

amount of formal engineering work and accurate within 10 - 15%.

Steps for a feasibility study

1. Geology and Resource: This is the step where drilling and sampling works is

performed. Various methods are available for drilling based on the soil and

mineral properties. The drill samples are prepared for the assay in order to

determine the minimum, maximum and average ore grade and these figures are

used to make the reserves estimation.

2. Mine design and Mineable Reserve: This is the step where most economic way of

mining is developed. Mine planning, model development, operation models and

cost analysis are performed and thus the mineable reserve is estimated based on

the economy.

The major steps for the mine development are:

mine access (surface/underground),

conveying system (especially in UG mines),

backfill requirement,

ore haulage, ventilation,

Selection of mining equipment and justified against the performance and

economy.

disposal of tailings generated.

3. Mineral processing facility: Sampling must be carried out to ensure that the

samples used in the mineral beneficiation processes are real representative of the

ore body. Some major characteristics of the ore body is determined prior to the

development of the plant design which includes Grinding work indices, feed size,

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settling characteristics, filtration characteristics etc.

Sometimes a mineral processing tests are performed in order to determine the

amenability of the given ore to different concentration technologies. The major

processes that are looked at are:

Crushing and grinding,

Concentration (Sizing, Gravity or Flotation)

Dewatering (Mechanical or filtering)

Chemical extraction (especially for gold)

When these tests are completed, based on the test results the basic material flow

sheet is developed. This helps in the selection of the equipment selection and the

stages of processing.

These data are used to estimate the amount and grade of concentrate, middling

and tailings that are used to search potential customers and revenue earned.

4. Tailings disposal: Tailing disposal system plays a crucial role in order to get the

mine permit. Mostly the tailings didn't place any major challenges. But, if the

tailings have hazardous or toxic materials like cyanide, mercury etc. in it, then the

disposal system must be effective in order to reduce the harmful effect on the

environment and society.

5. Infrastructure development: This section includes the civil and major earthworks

required to start the production. The office, labs, storage units, plant buildings,

mining equipment shelters etc. are included in the infrastructure.

6. Power supply: Determining the power source, power line distribution, total power

required and the power cost are the major things to be looked into in this step.

7. Water: Most of the plant processes are water based, so, the estimation of water

requirement plays an important role in the feasibility studies.

8. Environmental impacts: For a project to be permitted by any government, an

environmental clearance is required. In order to get the clearance, the

environmental impacts need to be studied. The important aspects are acid mine

drainage, cyanide management, and other toxic material controls (Arsenic,

mercury, sulfur etc.)

9. Other key parameters: Support facilities, maintenance, transport cost of man and

material, labor cost, site access (road facility or construction, fly in fly out,

marine etc.), social impacts are also need to be studied and the steps for social

responsibility.

10. Cost estimation: Based on the entire above-mentioned steps, capital and operating

cost for each unit is estimated. It included all the costs for mine equipment,

process equipment, construction costs etc.

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11. Financial Evaluation: This is the stage where the project is evaluated based on the

economy. The total cost and expenses are looked against the expected revenue

gained from the selling of final products and by-products.

The key financial indicators examined to determine the viability of the project

include Net Present Value (NPV) and the Internal Rate of Return (IRR). Annual

cash flow need to be estimated over the entire life of the project, from

construction to reclamation phase, based on clear and realistic capital

expenditures mine and mill operating costs, employee wages and sales revenue.

12. Sensitivity Analysis: A sensitivity analysis is then carried out to determine the

impact of variation in metal price, operating cost, metal recovery, metal grade,

and capital cost on the overall project NPV and IRR values.

The viability of the mine project is established by all these stages and if based on these

considerations if mine is feasible, then the next stage of actual development occurs.

Design elements of Underground Metal Mine (UMM)

The following constitutes the elements of underground metal mine design

1. Mineral resources and reserves i.e. mineral inventory

2. Cut-off grade

3. Production rate and mine life

4. Price of the mineral

Classification of Mineral resources

Of all the aspects of mining operations, the ore deposit and its characteristics is the only

aspect which is unalterable. Therefore the viability of a mining project is dependent on the

knowledge of mineral resource.

Geologists distinguish between mineral resources and reserves. The term resource refers to

hypothetical and speculative, undiscovered, sub-economic mineral deposits or an

undiscovered deposit of unknown economics. Reserves are concentrations of a usable mineral

or energy commodity, which can be economically and legally extracted at the time of

evaluation.

• Mineral resources is the name given to minerals which contain elements such as gold,

silver, copper, lead, zinc, iron, aluminum, nickel, molybdenum etc., as well as fossil

fuels, like oil, natural gas, and coal

• Mineral reserves are concentrations of various minerals and it is a geological term.

Whether a mineral deposit is also an ore deposit depends on its economic value.

• "Ore deposit" is therefore an economic term of a mineral deposit.

Mineral inventory (stock ) is commonly considered in terms of resource and reserve.

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Fig 1 Classification of Mineral Resources

Fig.2 Losses of various types in an u/g. metal mine

In terms of the mining project a mineral resource is divided into three categories as follows:

Geological resource (QG)

Mineable or workable reserves(QW)

Commercial reserves (QC)

INFERREDSUB-ECONOMIC

RESOURCES

DEMONSTRATEDSUB-ECONOMIC

RESOURCES

INFERREDMARGINALRESERVES

MARGINALRESERVES

INFERREDRESERVES

RESERVES

SPECULATIVE

HYPOTHETICAL

INDICATEDMEASURED

PROBABILITY RANGE

INFERRED

DEMONSTRATED

UNDISCOVERED RESOURCES

IDENTIFIED RESOURCES

Economic

Marginally

Economic

Sub-

Economic

Eco

no

mic

Fea

sib

ilit

yCertainty Of Existence

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Fig 2 . Reserve Classification

𝑄𝑊 = 𝑄𝐺 − 𝑄𝑁𝑊 (𝑄𝑁𝑊 = 𝑞𝑢𝑎𝑛𝑡𝑖𝑡𝑦 𝑜𝑓 𝑛𝑜𝑛 − 𝑤𝑜𝑟𝑘𝑎𝑏𝑙𝑒 𝑟𝑒𝑠𝑒𝑟𝑣𝑒𝑠)

𝑄𝐶 = 𝑄𝑊 − 𝑂𝐿 (𝑂𝐿= various unavoidable losses of ore reserve in

pillars, etc)

Cut-Off Grade:

Cutoff grade can be defined as the minimum grade of metal present in the mine which

could be mined economically. Cut-off Grade can be used to separate two courses of

action i.e. mine or to dump. The grade of mineralized material below cut-off grade is

classified as waste whereas the material above cutoff grade is classified as ore.

The cut-off grade is extremely crucial with respect to economical, production and

geological parameters of the mine. Too high a grade can reduce the mineral recovered

and possibly the life of the deposit whereas too low a cut-off would reduce the

average the average grade ( and hence profit) below an acceptable level.

Cut-off grade can be classified into two basic categories namely fixed cut-off grade

and the variable cut-off grade.

The fixed cut-off grade assumes a static cut-off for the life of the mine while the

variable cut-off grade assumes dynamic cut-off maximizing the mine present value.

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Professor Lane outlined three distinct stages in amine operation namely ore

generation (mining), concentration (milling), and refining.

The various factors which are essential for assessing cut-off grade for mining

operations are the type of ore resource/reserve present, extent of mine development or

present day cost development of mine, cost of drilling, mucking and transportation,

present value of revenues to be obtained from selling the ore, net cash flows have to

be considered.

For each of the stage as mentioned, there is grade at which cost of extracting the

recoverable metal equals the revenue from the metal. This is commonly known as

break-even grade. If the capacity of the operation of an operation is limited by one

stage only, the break-even grade for the stage will be the optimum cut-off grade.

Where an operation is constrained by more than one stage optimum cut-off grade may

not necessarily be beak-even grade. In such a case balancing the cut-off grade for

each pair of stages need to be considered as well.

Fig. Influence of cut-off grade on mining design parameters

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Fig. Optimum Mine Production rate

Categories of resources based on degree of assurance of occurrence

Identified (Mineral) Resource: Are the specific bodies of mineral-bearing material whose

location, quantity, and quality are known from specific measurements or estimates from

geological evidence. Identified resources include economic and sub-economic components.

To reflect degrees of geological assurance, identified resources can be divided into the

following categories:

Measured: Are the resources for which tonnage is computed from dimensions revealed in

outcrops, trenches, workings, and drill holes, and for which the grade is computed from the

results of detailed sampling. The sites for inspection, sampling, and measurement are spaced

so closely, and the geological character is so well defined, that size, shape, and mineral

content are well established.

Indicated: Are the resources for which tonnage and grade is computed from information

similar to that used for measured resources, but the sites for inspection, sampling, and

measurement are farther apart or are otherwise less adequately spaced. The degree of

assurance, although lower than for resources in the measured category, is high enough to

assume continuity between points of observation. Demonstrated: A collective term for the

sum of measured and indicated resources.

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Inferred: Are the resources for which quantitative estimates are based largely on broad

knowledge of the geological character of the deposit and for which there are few, if any,

samples or measurements. The estimates are based on an assumed continuity or repetition for

which there is geological evidence. This evidence may include comparison with deposits of

similar type. Bodies that are completely concealed may be included if there is specific

geological evidence of their presence.

Categories of resources based on economic considerations.

Economic: This term implies that, at the time of determination, profitable extraction or

production under defined investment assumptions has been established, analytically

demonstrated, or assumed with reasonable certainty (see guideline iii).

Sub-economic: This term refers to those resources which do not meet the criteria of

economic; sub-economic resources include Para-marginal and sub-marginal categories.

Para-marginal: That part of sub-economic resources which, at the time of determination,

almost satisfies the criteria for economic. The main characteristics of this category are

economic uncertainty and/or failure (albeit just) to meet the criteria which define economic.

Included are resources which could be produced given postulated changes in economic or

technologic factors.

Sub-marginal: That part of sub-economic resources that would require a substantially higher

commodity price or some major cost-reducing advance in technology, to render them

economic.

Some definition related to mineral resources:

• Ore is a naturally occurring, in-place, mineral aggregate containing one or more

valuable constituents that may be recovered at a profit under the existing techno-

economic indices. In metal mines, the amount of ore is usually expressed in tons

(metric ton =1000kg),

• Grade is a measurement of the metal content of ore.

• The grade of precious metal ore is usually measured in grams per tonne. The grade of

ore bearing other metals is usually a percentage (the weight for weight proportion of

metal in the ore).

• The grade of ore from a mine changes over time. Mining of a lower grade is likely to

incur (other things being equal) a higher cost per unit weight of extracted metal.

The most important factor in the profitability of a mine is usually the price of the

metal that it produces.

• Dilution is the result of mixing low-grade material with high-grade material during

material production, generally leading to an increase in tonnage and a decrease in

mean grade relative to original expectations.

Reserves of minerals are difficult to determine as the value and costs of extraction and

metallurgical treatment and transportation costs determine whether the resource are

potentially economic. Because of these uncertainties, mineral, mineral exploration is a

program that raises even more uncertainties.

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Lecture 3

3.0 Mine development

Opening a new mine is an expensive, time-intensive operation. Most mines must operate for

years to cover initial start-up costs, the period of capital investment for mine development

without any return on the investment is known as gestation period Mining is the process of

extracting valuable minerals from the earth. Mining involves a number of stages which occur

in a sequence. This sequence of stages is known as the mining sequence. The mining

sequence covers all aspects of mining, including: prospecting for ore bodies, analysis of the

profit potential of a proposed mine, extraction of the desired materials and, once a mine is

closed, the restoration of all lands used for mining to their original state.

3.1 Sequence of a mining enterprise

The mining sequence is divided into six stages. Each stage represents a certain period in the

life of a mineral deposit. The stages, ordered chronologically from earliest and following the

order in which they occur, include:

1. Exploration - gather data about potential mineral deposits and acquire the rights to

harvest those mineral deposits

2. Evaluation - determine which mineral deposit has the most profit potential

3. Mine Development - construction of a mine or mines

4. Production - operation of the mine or mines

5. Closure demolition of the mine or mines and rehabilitation of all lands used for

mining

Mine develop involves construction of various types of openings within the rock mass It is

therefore important to identify the importance of different types of mine openings on the

basis of their specific role in the entire term or life of the mine. Based on these criteria all the

mine openings are categorized into three types of openings, such as:

Main access to the deposit, which connects the surface and the ore body is also the

called the primary development opening.

Net-work of the openings like the levels, cross-cut, raise & winze, etc. – secondary

opening; which is the access to the stope

Source of the ore (stope) also termed the tertiary opening.

The role of primary opening is to provide an access to the deposit from the surface and

therefore the life of these openings is as much as the life of the mine. The secondary

openings are next important development openings in terms of the life. The life term of a

stope, the tertiary opening, is the shortest compared to any other opening of the mine.

The primary development is creation of a main access from the surface to underground, such

as shaft, incline, decline, adit etc., and any development which generates a network of

openings connecting the main access and the main production zone (stope) are called the

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secondary developmental works. For example, levels, raises & winzes, ore pass, cross-cuts,

ore chutes, u/g electrical sub-station & mechanical workshop, first aid room, etc., are

categorized as secondary development openings. A stope, which the place of main zone of

mine production comes under tertiary development

3.2 Stages of Mine Development

3.2.1 Primary Development – access to the deposit

Access to the ore deposit is first operation, which establishes the entry to the mine. For an

underground metal mine, the modes of entry to a deposit are: adit, incline, decline, a vertical

shaft, inclined shaft. Based on the geometry, strike & dip dimensions of the ore deposit, and

depth one or more combinations of different modes of access is decided. Once the deposit is

accessed, in order to commence the mine excavation of ore, various types of constructions

within the rock mass are needed for various engineering purposes. Some of these openings

are vertical, inclined, parallel to the strike and along the dip etc. The shape and the cross

section of the excavation depend primarily on the target production, purpose of the opening

(transportation, ventilation, water outflow, etc.,), nature & stability of the rocks type, the

period of service.

Permanent access and service openings, as shown in the above figure, are expected to

meet rigorous performance specifications over a time span approaching or exceeding the

duration of mining activity for the complete orebody. For example the service shaft must be

capable of supporting high speed operation of cages and skips continuously. Ventilation

shafts and airways must conduct air to and from stope blocks and service areas. Main haulage

drives must permit the safe, high speed operation of loaders, trucks, ore trains and personnel

transport vehicles. In these cases, the excavation are designed and equipped to tolerances

comparable with those on other areas of engineering practice. The mining requirement is to

ensure that the designed performance of the permanent openings can be maintained

throughout the mine life. The magnitudes of the mining induced perturbations at any point in

the rock medium surrounding and overlying an orebody are determined, in part, by the nature

and magnitude of the displacements induced by mining in the immediate vicinity of the

orebody.

3.2.1.1 Selection of a suitable access to the deposit

The decision of selecting the suitable access to the deposit, between a vertical shaft and an

incline is based on the following factors:

depth of ore deposit, size and shape of ore body,

surface topography,

geological condition of the ore and overlying rock mass ( it also includes the strength

condition of ore body as well as the surrounding rock type.

time for development,

method of mining (stoping)

cost and choice of material handling system.

21

Incline is not suitable for a deep seated ore body. Because with the increase in the depth of

ore body the haulage distance, at the required gradient, increases enormously and

proportionately the cost of material handling also increases. The cost of maintenance of the

inclined roadway increases. Though the rate of advance for incline/decline/drift are better

than sinking a shaft, with the advent of modern mechanized methods of shaft sinking can give

higher advance rates. Fully loaded ore trucks can travel up the incline and can travel straight

to ore dump. For shaft mine cars are to be loaded on a level via an ore pass and chute and

hauled to shaft. This system is not as flexible as trucks. However when a complete cost study

is made the use of inclines is never economical for deeper ore deposits.

22

Fig . A-E different modes of access to deposits

Fig. Cross-section of a service shaft

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23

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24

3.2 Secondary development

There are two categories of secondary development; first type is development in the nearest

proximity of the stope, like the stope access levels and the second type of development is

concerned to a stope or in-stope development. The in-stope development such as drill

headings and slot raises, horizontal and vertical openings for personnel access to stope, and

ore drawpoints from the stope. The life of drill headings, slot raises, draw points, sill & crown

is limited to life of the stoping. The openings, such as haulage levels and ore passes which are

developed near stress filed zone of a stope orebody rock. Their operation life approximates

that of adjacent stoping activity.

3.2.1 Levels and Level Interval

Level is an opening developed along the strike direction of an ore deposit and is driven with

zero to near zero (1 in 200) gradient. It is considered as the secondary mine development

operation of an underground metal mine, because it opens out the extent of mineralization

and thus a level offers a scope for a detailed evaluation of grade of the mineral deposit. Every

single underground mine developmental operation is a capital intensive and there is a

significant degree of risk, because any increase in the length of development openings could

augment high capital expenditures. In this respect mine development, involving levels and

their interval is an important operation. The levels also offer the service of transportation, for

men and material, from the shaft to the production site. Of the many factors influencing the

selection of a suitable level interval, the important factor is to facilitate quick disposal of

broken ore from the workings

3.2.1.1 Level intervals

Underground mining of ore deposits is necessarily worked with multiple levels. A level

interval is selected which lead to lowest overall mining cost for the mine development and

exploitation plan chosen. Number of factors affects these costs and some of them are

following:

• geological and natural conditions of the deposit and country rock

• method of mining

• development layout

• method of drivages of openings

• life of openings, mine life

• other financial considerations

The selection of optimum level interval is usually dependent on the development cost

(construction, supporting). Generally development cost increase with the number of main

levels required whereas exploitation cost as well as convenience of access for the miners

decrease with increasing number of levels. From the point of view of cost, a long interval

between levels is desirable. However in case of high grade ore deposits preclude higher level

intervals. The levels are placed at a closer interval to avoid missing high grade ore bodies.

25

Speed of stoping and character of ground are related factors. Levels interval should be such

that stopes are completed and abandoned within the time that they can be kept open without

undue maintenance cost. In order to determine optimum level interval calculations of

development and exploitation cost for different assumed level intervals are made and plotted

graphically and the lowest overall mining cost point gives the optimum point as shown in

figure below. The current trend with mechanized high production method is to have fewer

levels with large level intervals and supplemented by less cost sublevels as required by the

stoping method adopted.

Fig Determination of optimum interval between levels for a hypothetical multi-level mine

Fig. Sublevel Open Stope

Exploitation

Development

26

Fig. Stope developmental openings, ore draw points, slusher drifts.

3.3 Parameters considered in the design of stopes- tertiary openings

A stope, as shown below, is the site of ore production in an orebody. The set of stopes

generated during ore extraction usually constitutes the largest excavations formed during the

exploitation of the deposit. The stoping operation, that is, ore mobilization form it’s in situ

setting and its subsequent transportation from the mine void, forms the core of the mine

production process. In order that the stoping operations are safe it is essential to assess rock

performance within the orebody, and in the rock mass adjacent to the orebody. It ensures the

efficient geomechanical and economic performance of the individual stopes, and of the mine

as a whole. The size of stopes is large relative to all the other mine excavations. Therefore the

location, design and operational performance of other excavations connecting the stope and

the main access play a dominant role.

3.4 Raising Methods

3.4.1 Manual raising method

This is a simple and most common method adopted in majority of the metal mines.

The unit operations followed in the construction of a manual raise are:

drilling and blasting

mucking and transportation

erection / construction of a manual platform or also known as scaffold

The workers stand on a platform or scaffold made of timber planks supported in stulls

or iron bars fitted into the footwall. The clamps used for supporting the platform are made in

standard lengths out of old rails.

Drilling & Blasting: Jackhammers / stoppers are used for drilling either wedge pattern or burn

cut pattern holes of 32 mm diameter and 1.5m deep. Before each round is blasted the

platform is dismantled. Immediately after blasting, compressed air is forced to the working

faces to remove the fumes of blasting. In longer raises sometimes a blower with a flexible air

duct is installed. Access to the faces is by a ladder way.

27

Mucking & Transportation: The muck (ore if the raise in driven within the orebody, or a

waste rock if the raise is placed in foot-wall rock) based of ore or waste rock are trammed by

a mine car to the nearest grizzly.

Construction of a scaffold: The stoppers can reach a height of 2m and it facilitates the

construction of scaffold after every two rounds of drilling and blasting. The scaffold is

advanced regularly so as to maintain necessary head room at the face. The broken rock rolls

down by gravity. The scaffold is constructed by fixing steel bars into the holes drilled in the

side walls

Limitations: A simple but a very tedious method and has a limitation of comfortable raising

operations upto 15m. Careful checking and dressing down of the loose rock by skilled

workers before allowing workers to go up is essential At Jaduguda mine of UCIL where

this method of open raising was adopted for a number of stopes, the longest raise driven

was 90 m at 450 inclination.

Fig. Manual Raising method

3.4.1.1 Two compartment method

This method of raising is adopted for vertical or very steep raises only. After initial

excavation from the lower level on the direction of the raise for 2m the raise is divided into

two compartments and the follows a conventional driving methods

Raising with shallow holes is started by cutting out a recess at the bottom level, from which

subsequent operations are performed. Work is done from stage 1. After firing a round of

holes the stage rests on two or three stulls 2 temporarily set into holes made in the walls of

the raise. It consists of wooden planks laid over the stulls. Holes 3 are drilled from the stage

28

by means of stoppers. After the drilling is completed the drilling equipment and the tools are

removed from the face and the holes are charged with explosives. Before firing, the ladder

way 4 of the raise is covered by inclined wooden planks 5 which guide the broken rock away

into rock, while standing under protection of the stage. Then the timber sets are erected and

the working stage is transferred closer to the face. As the face advances, the ladder

compartment is extended and equipped with ladders. Rope ladder 7 connects the upper

segment with the working stage.

The raising cycle comprises the following operations:

inspection and dressing down of loose rocks,

timbering extending the ladder way,

construction of the working stage and drilling,

removing the working stage,

charging and firing of the blast holes, and

clearing the smoke.

One of the drawbacks of the method of raising by firing shallow holes is the need for

performing a number of subsidiary tasks (like building the stages and ladder ways, their

extension, and repairs, etc.).

Fig. Fig. Two compartment method

29

3.4.2 Mechanized Raising

Raising and winzing is one of the common development operations in underground metal

mines. These are vertical or sub-vertical connections between levels and are generally driven

from a lower level upward through a process called raising. An underground vertical opening

driven from an upper level downward is called a winze.

Raises with diameters of two to five metres and lengths up to several hundred metres are

often are developed either by manual and or mechanized methods, depending upon the size

and the extent of mechanization of a mine. The openings so created may be used as ore

passes, waste passes, or ventilation openings.

Earlier raising was done by manual method which was time consuming and hazardous.

Developments of raise climbers and raise boring machines have made the process faster and

safer.

The unit operations such as drilling blasting, mucking and erecting the support and surveying

for marking the centre line of a raise are done manually. The raising is done either dividing

the available area into two-compartments or a single chamber.

height of raising is limited specially by conventional and raise climbers ladder

climbing and making platform is hazardous in conventional method

potential hazard of rock falling

surveying is difficult

In mechanical raise climber most of these difficulties are avoided and the most popular to this

kind are:

1. Jora raising method

2. Alimak raise climber.

3. Raising by long hole drilling

4. Raise borers

3.4.2.1 Jora raising method

Jora raising method is suitable only for the condition when two levels are available for

connectivity by a raise. The method consists of drilling a large diameter hole at the centre of

the intended raise to get through into the lower level (Fig. below). From the upper level a

cage is suspended using a flexible steel rope that can be hoisted up and down using a winch.

There is a working cabin also known as Jora cabin. The Jora cabin is provided with a sturdy

working platform on top of it, it is from this platform that the drill operators make the drill

holes.

Drilling: Usual practice is to follow parallel hole pattern and the central hole is used as a

relief hole. A stopper is used for drilling the holes of 34 mm diameter. Before blasting the

entire jora cabin is lowered to the lower level.

30

Limitations:

1. One of the main limitations is that two levels are essential and arrangements are made

in both the levels.

2. The need to drill large diameter central hole for the hoisting rope.

3. Slow and a tedious operation.

4. Rate of advance is low.

1- Winch for rope; 2- winch skid; 3- drilling platform; 4- hoist rope;

5- Jora cabin; 6- steel rope; 7- Hole reel; 8- Drill hole for steel rope

3.4.2.2 Raising by Large Diameter Blast Holes

Top level

Bottom level

31

Fig. Raising by Large Diameter Blast Holes

3.4.2.3 Alimak Raising:

Alimak raising is a mechanised blind raising method. It was introduced in mines way

back in 1957 and over the time it has proved to be economical, flexible, and a safe method of

raising for as long as 900 m. It can be used for vertical and inclined raises.

The machine along with a cage runs up and down on a guide rail that incorporates

rack and pinion gear mechanism (Fig. below). The guide rails are in segments and fastened to

the rock by rock bolts. They are extended as the raise advances.

The drilling operation is carried out standing on the platform after charging the holes

the cage is taken down at to a safe place for blasting the face. After the fumes clearance the

cage goes up again and guide rail extension is done. The blasted muck is removed.

Fig. Rack-and-pinion gear mechanism

32

Alimak raising provides the safest of all entry methods involving the least risk to the

miner and can excavate safely through all types of ground conditions supporting the face after

each blast is taken ensuring the integrity of the excavation during all stages of development.

The Alimak raising system ensures fast mobilisation, minimal preparation, is flexible,

accurate, economical and very cost effective even over short distances. Even multiple raises

with directional changes in the raise of up to 90° can be carried out easily making this method

the ideal choice for ore passes, crusher chambers, split level ventilation raises or any difficult

excavation profile.

Alimak raise climbers are widely being used to drive shafts and raises in Mount Isa

mine Australia. Importantly the longest Alimak raise developed to date in these mines is

more than 1000m in length.

Fig. Preparatory work for installation of Alimak raise climber

Cycle of Operation

Step -1(Fig. a) –Drilling; Drilling is undertaken from the drill deck on top of the raise

climber, which is sized to suit the size, shape and angle of the raise. Drill machine is jack

hammer for drilling a 34 mm diameter and 2 m long blast holes. Burn-cut parallel blasting

patter in the common pattern used for raise blasting.

Step -2 (Fig b)-Loading: When drilling is completed the face is charged with explosives

along with MSD & HSD delay detonators. Of all the rounds, perimeter round is very

33

important in raise blasting, and smooth blasting techniques are followed to contain over-

break.

Step-3 (Fig c)- The Alimak climber is then lowered to the bottom of the raise and into a

station for protection before the blast is triggered from a safe location.

Step-4: Ventilation: The Alimak system provides for efficient post blast ventilation and a

powerful air/water blast effectively dislodging loose rock from the freshly blasted face

making ready for re-entry.

Figure Steps of operation in Alimak raising method.

34

This method has the following advantages:

permits driving of long raises

personal are well protected in a cage under the platform

the miners work from the platform that can be easily adjusted for convenient height

timbering is avoided and stability can be increased by rock bolting if necessary

no danger from falling of rock pieces

However the cost and other arrangements required cannot justify this for short raises. Figure

above shows complete cycle of raising.

Special feature of Alimak raise climbers:

A. Drive Units:

The raise climber is developed with three kinds of drive units: air driven, electrically driven,

and diesel/hydraulically driven.

Of the different types of Alimak raise climbers, compressed air driven raising is very

common in the country, followed by diesel operated raise climbers are popular.

Air Driven:

In the air driven raise climbers, compressed air comes through a hose. An automatic winch or

reel winds the hose up and down as per the movement of the alimak in the raise construction.

The air motors are effective for raising up to 200m length.

Electrical drive:

Electric are not common in mines, however they have a capacity of driving about 1000m long

raises. The longest vertical raise for ventilation shaft at the Densison mines, Ontario, Canada,

in 1974 [SME-UMM Hand book].

Diesel / Hydraulic drive:

Diesel operated Alimak raises climbers are also common after the compressed air driven

machines. However there is a risk of excess air pollution due to diesel operated machines

underground. The diesel/hydraul;ic driven raise climber can drive more than 1000 m long

raises in one step.

35

The figure above gives the scope and limitation of various types of Alimak raise climbers.

B. Safety features

For the types of Alimak raise climbers the following safety features make them more

adoptable in mines;

Over speed control system; the permitted speed limits on descent are 0.9m/s, if the

climber exceeds this speed limit the automatic braking system stops the climber to

further descend.

The rack-and-pinion gear plates are wielded to the guide rails thus ensure a guided

manoeuvring of the climber up and down the raise.

The cross section of a guide rail is as shown in the figure below

(a) (b)

Fig (a). Cross-section of a guide rail; (b). Rack-and-pinion mechanism

36

The air, water supply is provided through the ports within the guide rail,

approximately 25m3/min air supply is provided continuously at the face point. This

facilitates the operators with fresh air at the working face. There is a provision to

increase the air quantity as per the requirement.

Telephone communication between the face crew and the bottom crew is provided by

an insulated wire passing through one of the ports in the rail.

Blasting cable also runs through the port within the rail.

A canopy is also provided for the safety of the face workers while scaling down the

loose material from the roof.

C. Initial guide rail sections

The guide rails for negotiating the curves are special made in angular sections, 80, 250, 250,

250, 80 and having a radius of 2.3 ~3 m for vertical raises. The brow point is the point where

the cross cuts terminates into a vertical raise (Fig below), is slashed at 450 to accommodate

the circular guide rail segments.

RAISE BORING METHODS

Raise-Boring

In this system, the pilot hole is drilled down to a lower level in the mine or civil project. Once

the pilot hole connects to the lower access level in the rock, the drill bit is removed and a

reamer or raise head is attached and the reamer is rotated and pulled upwards. The broken

rock falls to the lower level by gravity. This system operates with the drill string in tension

and this provides the most stable platform.

Brow

37

Figure. Raise Boring

Down-Reaming

In this system, the pilot hole is drilled downwards until it connects to a lower access level.

The drill string (all drill rods, stabilizers and cutting bits) is retrieved and then a reamer is

pushed downwards. The cuttings flow down the previously drilled pilot hole. This method

uses drill string in compression and usually stabilizers must be installed to eliminate the

potential of the drill string buckling.

Figure: Down Reaming method of raise boring

38

Box-Holing

The most difficult raise method, known as Box-Hole excavation. It is to drill a pilot hole to

any level up from the raise borer. Once the desired length is achieved the drill string is

retrieved, and a reamer attached and pushed upwards. The broken rock falls down the

enlarged hole onto a special collection chute attached to the top of the raise borer. This

technique has been largely used to replace ladder rises, which completes the box-hole using

conventional methods. Ladder rise excavation is very dangerous

.

Figure. Box-holing method of raising.

ADVANTAGES OF BORED RAISES

Raise boring offers several advantages over the conventional drill and blast method.

The most important are safety, speed, physical characteristics of the completed hole,

labour reduction and cost reduction. The safety factor in raise drilling cannot be over

emphasized. No men are exposed to the danger of rock fall from freshly blasted

ground or to the continual use of explosives, with their fumes and inherent danger of

misfires. Raises can be safely drilled in ground that would be extremely hazardous, if

not impossible, to drive by conventional methods.

A hole drilled by Raise Boring Machine can generally be completed in a fraction of

the time required for conventional methods. The bored raise, with its firm undisturbed

walls, is more adaptable to use as ventilation and rock passes. As conventional

methods require a relatively large opening, it has become customary to drive raises

larger than actually required for ore and rock passes, a fact that long experience has

borne out. The advantage of smooth walls in ventilation raises is well known.

39

Raise boring will not only reduce labour requirements by achieving a higher advance

per day but, along with another technological advances, will have the tendency to

attract a higher level of skilled labour to the mining industry.

Last, and probably most important from the long-range viewpoint, is cost reduction.

Although, it is true that the direct cost of conventional raises, especially short ones,

may currently be less in many cases, labour and material costs are continually

escalating and therefore their costs increasing. Skilled conventional miners, always in

short supply, are not required to operate a Raise Boring machine. Improved raise

drills, drilling techniques, pilot bit and cutters are lowering the cost of machine

excavated (RBM) raises. Less total manpower, less rock to handle, less construction

time and increased safety all add up to less costs and earlier projects.

Shaft Station

Underground mining operations involve deployment of different types of heavy duty rock

excavation and transportation machines. Some are electric power driven, others are diesel

operating machines. There are a few specialized openings such as bunkers, pumping station,

electric sub-station etc., at the bottom of the main shaft, and it is the horizon where the

vertical shaft intersects with horizontal openings. This is known as the shaft station.

The shaft station serves as the principal terminus of all underground and surface operations.

Those related to materials handling involve: skip loading pockets, retention bunker;

ventilation arrangements; pumping stations; electrical sub-stations; underground mechanical

shop / workshop; first aid centre & rest rooms etc.

The design considerations depend on the number of shafts within the station, type of deposit,

mode of materials handling in the mine and in the shaft, water inflow, ventilation

requirements, mining equipment, etc.

Fig. Standard shaft station layouts

40

a-with circular mine traffic; b- with shuttle traffic; c- loop like layout of shaft staion;

1- Main shaft; 2- service shaft

Shaft station is an aggregate of working located in the immediate vicinity of the shaft. These

are provided to afford connection between a shaft and the different levels in a mine. Their

primary use is to tenable men and material to be delivered at the different working horizons

and for raising the ore. The size of the station will depend on the size and amount of material

that it will be required to accommodate.

Generally the longer the life of a mine and larger the output the shaft station becomes more

complex. Some of the factors that are considered for design of shaft station are:

Type of deposit

Mode of material handling in the mine

Hoisting of ore in the shaft

Water inflow and ventilation

Mining equipment

Shaft stations related to the material handling are skip loading pockets, retention bunkers

pump chamber, explosive storage chamber, locomotive room and sometimes primary

underground crusher. These chambers are important link in the extraction process, transport

etc. They are located near the main or auxiliary shaft because of their functions.

The first group of chambers includes explosive storage, pump house, miners’ rest room

where as locomotive repair and clearing, dispatcher rooms are related to the transport. The

construction of shaft station chamber is made by conventional drilling and blasting method

taking into consideration of ground conditions. These chambers are properly supported by

bolting, grouting etc.

Question

Explain with a neat sketch a shaft with skip hoisting system for a production level of say,

1200 tpd . Show the surge bin, loading pocket, measuring hopper excavated and installed in

the shaft station label the sketch ?

Answer

The shaft stations in hard- rock mines for material handling arrangement will have the

following:

1. Skip loading pockets,

2. Retention bunkers

3. Pump chambers

4. Main power station

5. Explosive storage chamber

6. Locomotive room

7. Mechanical & electrical workshop

8. Dump (ore/waste) chamber – with bunker & u/g crusher.

41

9. Arrangements for the type of ore/waste transport system ( eg: belt; train)

1- Access drift to waiting room; 2- basement for two-level traffic and swinging platforms;

3- Basements for pushers and barrages (blocking cars); 4- a slot for control equipment

Fig. Inset of cage shaft with three levels to step in and out for crew.

The size of the inset of a cage shaft depends on the width and number of cages being hoisted

on this level, number of decks in cages, and length of the supplies to be delivered. Depending

on the skip loading system and horizontal transportation arrangements, there could be the

following sets of openings for loading facilities:

1. For rail transport :

a. Dump(tippler) chamber or unloading ramp (for Granby cars), batchers chambers( this

for accommodating a batch or a train of mine cars), skip chamber

1-Skip chamber; 2- batcher chamber; 3.- tippler chamber; 4- basement of shifting mechanism; 5- basement of

braking system; 6- drive slot; 7- electrical equipment slot; 8- ventilation slot.

Fig. Connection of production skip shaft with the opening of loading system for rail transport system.

42

b. Dump(tippler) chamber or unloading ramp (for Granby cars), retaining bunker, loading

devices chamber, batchers chambers( this for accommodating a batch or a train of mine

cars), skip chamber

1- Skip shaft; 2- skip chamber; 3- batchers chamber; 4- switches chamber 5- loading chamber; 6- retaining bunker; 7-

distribution chamber; 8- distribution ramp; 9- drift for clearing away jams; 10- chute

Fig Connection of production skip shaft with the openings of the loading devices for horizontal rail transport.

c. For belt transport: unloading chamber, retaining bunker, loading chamber, batchers

chambers( this for accommodating a batch or a train of mine cars), skip chamber

1. Skip shaft; 2- skip chamber; 3- belt scale ; 4- retaining bunker; 5- unloading chamber.

Fig. Connection of production skip shaft with the opening of loading devices for horizontal belt

transport system.

43

Lecture 4- Stope Development

Once the economic extraction of ore body is ascertained, the step follows next is

development and preparation stope for extraction or ore. The development of an ore drift

(cross-cut) will confirm the thickness (extent of orebody) and continuity of the ore body and

enable the planners to finalize stope design.

Different development configurations and construction arrangements are possible for ore

body geometry. The stope preparation involves development of haulage level and sill-level.

This approach allows the development of draw points (figure below)

Fig Plan view of development of ore and footwall drives.

Draw points are developed at the bottom of open stopes as an inverted cone by drilling and

blasting. Their form is determined by the way in which the ore is to be loaded.

A large chute can be used to load ore from a main ore pass into a dump truck or smaller

chutes can be installed on each of several ore passes along a level to load directly into mine

cars.

Figure shows ore loading chutes. Chutes cause production holdups if they become blocked by

large pieces and to exclude the large pieces from coming to chute, ore is fed through grizzly

which has a grating made up of steel bars. Lumps which do not fall through grizzly are

broken with hammer of pneumatic pick.

Fig. Ore loading chutes

44

The figure below shows a typical draw point configuration for LHD/Shovel loading draw

point. In this configuration the draw points are usually 10m long and driven perpendicular to

the haulage-way to facilitate ore loading into mine cars. The interval of draw points is around

10m apart. The dimensions of these draw points are selected considering the ease of loading.

The draw point around the mouth or the entrance of the stope requires a lower back to

establish a brow that will prevent ore from spreading too far into the draw point.

Fig. LHD/ Rocker shovel draw points

T

Plan view of the draw point with track system of transportation

45

Fig. Cross section of a draw point configuration-track system of transportation

Another form if scram (also known as scraper) driven draw point. Ore is broken in the stope

and gravitates down into the drive. A scraper bucket is used in the drive to scrape ore and

drop it down through a grizzly down a mil hole into mine cars. Figure shows a scram driven

draw points and mill holes. Another from is to load ore from a stope by a mucking machine,

figure showing LHD draw points.

Fig. Scram drive points and ore draw points

In some mines construction of individual draw points for open stopes in not carried out. The

stope bottom is percussive drilled from the draw point level and blasted into a continuous v-

shape. Broken ore is loaded out from the bottom drive as it comes down. It is still necessary

to drive a raise to form an initial cut-off slot. Figure shows v-shaped draw point. A sill pillar

is left horizontally around and above the level drive to protect them and provide height to

develop draw points. As stopes are worked upwards to meet the level above a horizontal

crown pillar is left below the level above to stope them from collapsing.

Stope development thus includes haulage drifts cross cuts drifts, chutes and draw points,

raises. The size of the development is dependent on the equipment and winning methods to

be used. Minimum development requirements for a typical ore body include a drift from the

46

main haulage to the ore body, raising into the ore body, driving the stope sill and finally

installing draw points and chutes.

Fig Draw point

47

Fig . Mechanised ore loading methods

Ore pass system

Ore passes are underground passageways for the gravity transport of broken ore, waste rock

from one level of a mine to a lower level. Inclination of ore pass varies widely within a range

of 450-900, and most common angles are 700 and cross sections are mostly circular. Besides

transport of ore it also sometimes serves as a storage which is required for efficient mines

operation. Ore pass length range from 10 m to 200m or more

The components of ore pass system include: (1). a raise connecting two or more levels, (2).

Top-end facilities for material size and volume control such as grizzles, crusher and (3).

bottom end structures to control material flow.

Unlined ore pass may be located in country rock (FW) but some mines are lining ore-passes

with steel fibred-reinforced shotcrete. The bottom of the ore-passes at the haulage level

usually contains a loading chute equipped with pneumatic / hydraulic operated gates. The ore

is loaded in to tubs and a train of tubs then dump the ore in the main ore-pass which is usually

located at a haulage shaft.

48

Fig. Schematic of an ore-pass: tip section; discharge zones.

In mechanized stopes the ore is removed from the stope by LHD units and is dumped at the

stope ore pass for handling at the lower level from where it is transported and dumped in the

main ore pass. The main ore pass are developed within the ore body rock or within the ore

body peripheral rock. Their operational life approximates that of adjacent stoping activity and

in some cases the excavations may be consumed in the stoping process.

Proper design of ore pass requires that the broken ore, waste rock will flow when the outlet is

activated. The flow process is driven by gravity and resisted by friction and cohesion. Proper

design will see that their malfunctions of ore pass operations are to be prevented: failure to

flow resulting in hang-ups and failure to flow over the entire cross-section of the ore pass

referred to as piping. The other important design consideration is the stability of ore pass

walls.

Ore pass construction

Ore pass systems are an integral part of the materials handling system in the majority of

underground mines. Ore passes are developed using either mechanical (raise borer) or drill

and blast techniques (Alimak, conventional raising and drop raising). The conventional

manual method of raising is slowly being replaced by Alimak raising. In Quebec mines,

Alimak raising was used in 63% of driven ore passes while only 3% were raise bored. The

dominance of Alimak driven passes over raise bored passes in Quebec mines is attributable to

several causes. It ensures a reasonable degree of safety for the miners, while still allowing the

installation of support. Furthermore, the ability to drive the Alimak pass from a single access

49

(as opposed to raise boring, which requires that both the bottom and top accesses be

developed) and a strong expertise of local mining contractors are also contributing

factors.Conventional and drop raises represent 29% and 5% of the sections, respectively (Ref:

Ore pass practice in Canadian mines by J. Hadjigeorgiou, J.F. Lessard*, and F. Mercier-Langevin; The

Journal of The South African Institute of Mining and Metallurgy vol. 105 Dec. 2005). The dominance of

Alimak raising is attributed to several reasons. It ensures a reasonable degree of safety for the

miners, while still allowing the installation of support. Furthermore, the ability to drive the

Alimak in blind raises (as opposed to raise boring, which requires that both the bottom and

top accesses be developed) and it provides comfortable working environment at the face.

Table Case example of U/G mines of Lead & Zinc Quebec, Canada

(Ref: Ore pass practice in Canadian mines by J. Hadjigeorgiou, J.F. Lessard*, and F. Mercier-Langevin; The Journal of The South African

Institute of Mining and Metallurgy vol. 105 Dec. 2005).

Ore pass section length

50

There is an inherent relationship between the type of excavation method and section length.

Typically, sections excavated by drop raising or conventional rising are shorter than sections

driven by Alimak or raise borers.

There are several practical and financial considerations that influence the selection of an ore

pass length. If, for example, an operation aims to minimize its capitalized development, it

will end up driving short ore pass sections, going from one level or sub-level to the next,

concurrently as the various levels are entering into production. Quite often a mine that

experienced problems when driving and operating long sections will subsequently opt for

shorter sections when constructing new ore and waste passes. An excavation of greater length

is more likely to intersect zones of poor ground. It also has a higher potential for problems

and is harder to bypass. Longer sections may also result in higher material flow velocity in

passes operated as flow-through.

Ore pass section inclination

Ore pass inclination varies between 45° and 90°, with an average inclination of 70°. The

choice for a particular inclination is dictated by the need to facilitate material flow. Shallow

sections may restrict flow, especially if a high proportion of fine material is present, while

steeper excavations result in higher material velocities and compaction. It should be noted

that all vertical sections are shorter than 100 m. Generally steep ore passes (80º ± 8.3º) are

advantageous because it ensures continuous material flow and limit hang-up occurrences.

Ore pass section shape

The majority of excavated ore passes are square or rectangular. Circular sections are usually

associated with raise boring methods but in some instances, circular sections were excavated

using Alimak. In most cases, the main factor indicating the choice between a rectangular and

a square section is local mine experience. Circular shape was used based on anticipated

higher stress regimes. It is of interest to note that a review of ore pass surveys reveals that

under high stress, and with material flowing in an ore pass, a design circular shape is not

maintained for long (in unlined ore passes). Ore pass size is an important factor influencing

material flow. This is reflected in empirical guidelines linking the potential for hang-ups with

ore pass size and material size. A common dimension of 2.0 m is widely used, however there

are some mines where a relatively larger cross-sectional dimension of 2.5 ± 0.6 m have also

been adopted.

Finger raises

Finger raises are used to funnel material into a pass intersecting two or more production

levels. Typically, a finger raise is a square opening with a smaller cross-sectional area than

the rock pass it feeds. The most common dimensions for a finger raise are 1.5 and 1.8 m.

51

Screening of oversize material

Oversize material dumped into the passes may lead to blockages or interlocking hang-ups.

This can be avoided by either instructing the mucking crew or by installing the necessary

infrastructure to restrict the entrance of the oversize material.

The mechanical method of retaining oversized material at the mount of an ore-pass is by the

installation of a grizzly. Sometimes mucking crews can be ‘persuasive’ in trying to push the

block through the bars with the bucket. This practice damages both the bars and the scoop.

Broken and missing bars are often the result of this practice. In addition, the intrusion of a bar

in the ore pass can lead to severe obstruction further down the system. Grizzlies are the best

to keep big blocks out of the passes. Grizzlies require less maintenance than scalpers.

Reinforcement

Resin-grouted rebar constitutes the most popular reinforcement type for ore pass systems.

Nevertheless, the most recently developed excavations are reinforced by resin grouted short

cable bolts. An ore pass section is considered to have ‘failed’ if it had expanded to twice its

initial volume as recorded in the original layout.

Ore pass problems

Analysing the causes of degradation is a complex process due to the potential interaction of

several mechanisms. There is a relationship between the material unit weight and the degree

of observed degradation of the walls of the ore pass. A qualitative assessment of the dominant

degradation mechanisms include: structural failures facilitated by material flow; scaling of

walls due to high stresses; wear due to impact loading caused by material flow; wear due to

abrasion and blast damage caused by the hang-ups clearing methods.

Wall damage attributed to impact loading is most often localized at the intersection of finger

raises to the ore pass. It is most probable that the presence of structural defects in the rock

mass accentuates the influence of impact loading, resulting in more pronounced degradation.

The use of ‘rock boxes’ can reduce impact damage but in most cases impact damage is

localized on the ore pass wall facing the finger raise. Abrasion rate depends on the

abrasiveness of the material and the ore pass walls’ resistance to abrasion.

Blockages

Blockages are the most commonly encountered type of flow disruption in ore pass systems.

Flow disruption near the chute may be due to blocks wedged at the restriction caused by the

chute throat. Another source of problems is caused by the accumulation of fine or ‘sticky’

material in or near the chute, on the ore pass floor. This reduces the effective cross-sectional

area and results in further blockages.

Material flow problems

52

Some types of material flow problems are reported in every mine operating an ore pass

system. Sometimes the transfer of coarse material can result in hang-ups due to interlocking

arches, while the transfer of fine material results in hang-ups due to cohesive arches,

Hang-ups

Restoring material flow is a priority in operating mines. There are several methods to restore

the material flow in case of a material hang-up with in the ore pass and they can be classified

as those that employ water and those that rely on explosives,

Most hang-ups lower than 20 m are brought down by attaching explosive charges on wood or

aluminium poles used to push the charge up to the hang-up. As a last resort, holes drilled

toward the hang-up can be driven and explosive charges set inside the hole, near the supposed

hang-up location. If the location of the hang-up is not clearly identified, it may take more

than one attempt to restore flow.

Cohesive hang ups are difficult to dislodge using explosives. Some operations resort to

blowing compressed air through a PVC pipe raised up to the hang-up location or dumping a

predetermined amount of water from a point above the hang-up. All mines have strict

procedures about the use of water in order to avoid the risks of mud rushes.

Fig. Hang-ups in an ore pass due to (a) interlocking; (b). cohesion arching,

53

Fig. . Damage zones in an ore pass.

ORE PASS DEGRADATION DUE TO IMPACT

(ref: Influence of finger configuration on degradation of ore pass walls K. Esmaieli Université Laval, Quebec City, Canada J. Hadjigeorgiou University of Toronto, Toronto, Canada; ROCKENG09: Proceedings of the 3rd CANUS Rock Mechanics Symposium, Toronto, May 2009 ;

Ed: M.Diederichs and G. Grasselli)

In ore pass systems gravity movement of rock includes rolling, sliding and inter fragment

collision. The interaction of moving material and ore pass walls can result in the development

of wear and/or impact damage zones. Wear is associated with the particles rolling and sliding

along a surface resulting in the scouring of the wall surface. Damage attributed to impact

loads can be caused by single falling boulders in the ore pass, a stream of rock or a large mass

of material, Iverson et al. (2003). The mechanical properties of the rock mass along the ore

pass wall can influence the extent of damage. Stacey & Swart (1997) note that wear of ore

pass walls is greater in weak rock material and in the presence of stress scaling. If the ore

pass is located in a rock mass with structural defects the action of moving material can

initiate further wall degradation, including falls of ground. Ore pass wall damage, induced by

impact, is one of the most important mechanisms of ore pass degradation. This paper reports

on-going work, using numerical models, on the influence of material impact for several ore

pass and finger raise configurations.

Figure above illustrates a typical finger raise - ore pass configuration. Hadjigeorgiou et al.

(2005) report that, in Canadian underground mines, finger raises have cross section

dimensions of 1.5 m x 1.5 m and 1.8 m x 1.8 m. The fingers are linked to ore passes of larger

cross section dimensions. A well designed finger raise can minimize the ore pass wall

damage and maximize ore pass longevity. Current practice is often based on empirical rules

which quite general and may not always be appropriate for site specific conditions. Empirical

guidelines recommend an inclination of 60o for finger raises in order to ensure free flow of

rock fragments in the finger raise. This recommendation may not be valid for all the

54

conditions. The finger raise inclination influences the motion and interaction of rock

fragments flowing in the ore pass and the resulting load on the ore pass wall. If the finger

raises are steep this will result in higher impact velocity on the ore pass walls. On the other

hand if the finger inclination is shallow material flow is slow and can result in hang-ups. A

steeply inclined finger raise results in narrower pillars at the intersection of the ore pass and

finger raise which are more susceptible to stability problems. Consequently an operational

design will use a finger raise inclination that will minimize impact load on the ore pass wall

while maintaining material flow in the finger.

It has been demonstrated that particle impact velocity and kinetic energy increase with finger

raise inclination. The impact duration decrease with increase of finger inclination. These

observations can be used to evaluate different options of finger inclination for any particular

ore pass inclination. The analysis clearly demonstrated that the choice of intersection angle

has a significant influence on the resulting impact loads on the ore pass wall and the location

and magnitude of damage to the ore pass. The highest impact loads were reported for

intersection angles of 1400 and 1450.

Q. Explain the gravity ore transportation methods in u/g metal mines

Fig. Ore pass system in Mount Isa Copper Mines –Australia (Ref.L.J.Thomas Intro. to mining)

55

Lecture 5 Factors influencing the selection of a suitable stoping method

The following factors are considered in selecting a suitable method of stoping operation.

1. Mining excavations and their importance in terms of the life term of a mine

2. Rock mass response to stoping activity

3. Spatial distribution of the ore-body

4. Disposition and orientation

5. Size

6. Geomechanical setting

7. Ore body value and spatial distribution of value

8. Engineering environment.

1. Mining excavations and their importance in terms of the life term of a mine

The three types of openings are employed in the mine operation, these are the ore sources, or

stopes, the stope access pathways, or the levels, cross cuts; and the main mine service

openings – shafts, inclines, declines, or adits. The geomechanical performance of these

different types of openings is specific to the function of the opening. Based on their function

and the life term of these openings, they are categorized as:

Primary openings - shafts, inclines, declines, or adits, these are the permanent

openings in comparison to the other two types

Secondary – levels, cross cuts, raises & winzes, drifts, etc., - these are semi-

permanent openings, their life terms is relatively less compared to the primary

openings.

Tertiary openings: stopes or the source of ore – the main production zone. The life

term of the stopes is the shortest of the three above openings.

Stopes:

A mine has a large number of stopes therefore; a set of stopes constitutes the largest

excavation underground. The stability of stopes is controlled not only by the orebody strength

condition but also on the strength of the peripheral rock (HW and FW) the principles of stope

layout and design are integrated with the set of engineering concepts (like the rock

mechanics) and physical operations (such as mine transportation of the ore and waste) which

together compose the mining method for an orebody.

It is a commonly held belief amongst underground mine planning and design engineers

that in a sub-level open stoping mine, the bigger the stopes – up to the geotechnical limits –

the greater will be the production rate and hence, the more cost efficient the mine. This paper

shows that this can be a fallacy – it is usually true for the individual stope but may not be true

for the mine when considered as a system of inter-related stopes.

In a fixed size orebody there is a limit in the production rate achievable which in turn is

related to the number of active stopes, in the sense that the stopes are in some phase of the

stope development cycle (preparation, production, filling or curing) at a given time frame.

Once this limit is reached, there are no more stopes that can be brought into production. This

is a physical constraint, which places a limit on the production rate achievable for the stoping

56

system. However, this constraint, the number of stopes, can be changed. This can be

accomplished by either altering stope size or cut-off grade.

Fig. Division of the ore body into active workable stopes based on grade value

Fig. Longitudinal section of a mine

57

58

2. Rock mass response to stoping activity

The extraction of mineral resources involves rock excavations of different shapes, sizes, and

orientation based on the purpose for which the excavation is made. And it is obvious on the

creation of an opening (stope / drive) the state of equilibrium in the surrounding rock is

disturbed and the redistribution of the induced stresses is dependent on the type of rock mass,

size of the opening and method of excavation.

The dimensions of ore bodies of mining significance typically exceed hundreds of meters in

at least two dimensions. During excavation of an orebody, the spans of the individual stope

excavations may be of the same order of magnitude as the orebody dimensions. The

performance of the host rock mass during mining activity can be easily measured in terms of

the displacements of orebody peripheral rock. It is clear from the studies of stresses around

mine openings, the zone of influence is usually taken as 3dm, where dm is the minimum

dimension of the opening. The zone of influence is considered as the near field zone and the

zone outside this is termed the far field zone.

The rock mass response to stoping operations is dependent on the inherent strength of the

rock. Therefore on the basis of its response, a rock mass can be categorised into a class of

competent (strong and self-supporting) and in-competent (weak and crushing & crumbling

type of rocks). There are many rock types which fall in between these two extremes.

Therefore there can be stoping methods which are self-supporting, and a few stoping methods

need some artificial supporting and lastly there can be some which cannot be supported, such

stopes are left to crumble and cave down.

Fig. Rock mass response to mining

The supported methods of working can succeed only if the induced stresses are less

than the strength of the near-field rock. Caving methods can proceed where low states

Underground mining methods

Pillar supported Artificially supported Unsupported

Room & Pillar

SublevelLong holeOpenstoping

Cut-and-Fill Shrinkage VCR Sub LevelCaving

Blockcaving

Magnitude of displacement in country rock

Strain Energy storage in near-field rock

Rock mass response to Mining

59

of stress in the near field can induce discontinuous behaviour of both the orebody and

overlying country rock, by progressive displacement in the medium.

In supported methods, since the strength of the rock mass in higher, they exhibit the

ability to store more strain energy in comparison to the caving methods.

For caving method prevents the accumulation of strain energy by continuous

dissipation of pre-mining energy by fracturing.

Fully supported stopes may completely depend on natural support in the initial

stoping phase, using ore body remnants as pillar elements. In the early stages of pillar

recovery, various types of artificial support may be placed in the mined voids, to

control local and regional rock mass displacements. In the final stages of pillar

recovery, pillar wrecking and ore extraction may be accompanied by complete failure

of the adjacent country rock. This change in the state from one geomechanical basis to

another can have important consequences on the performance of permanent openings

and other components of a mine structure. This indicates that the key elements of a

complete mining strategy for an orebody should be established before any significant

and irrevocable commitments are made in the pre-production development of an

orebody.

3. Spatial distribution of the ore-body

This property defines the relative dimensions and shape of an orebody. It is related to the

deposit’s geological origin. Ore bodies described as seam, placer or stratiform (strata-bound)

deposits are of sedimentary origin and always extensive in two dimensions. Veins, lenses and

lodes are also generally extensive in two dimensions, and usually formed by hydrothermal

emplacement or metamorphic processes. In massive deposits, the shape of the orebody

is more regular, with no geologically imposed major and minor dimensions. Porphyry

copper ore bodies typify this category. Both the orebody configuration and its related

geological origin influence rock mass response to mining, most obviously by direct

geometric effects. Other effects, such as depositionally associated rock structure, local

alteration of country rock, and the nature of orebody–country rock contacts, may impose

particular modes of rock mass behaviour.

4. Disposition and orientation

These issues are concerned with the purely geometric properties of an ore body, such as its

depth below ground surface, its dip and its conformation. Conformation describes orebody

shape and continuity, determined by the deposit’s post-emplacement history, such as episodes

of faulting and folding. For example, methods suitable for mining in a heavily faulted

60

environment may require a capacity for flexibility and selectivity in stoping, to accommodate

sharp changes in the spatial distribution of ore.

5. Size

Both the absolute and relative dimensions of an ore body are important in determining an

appropriate stoping method. A large, geometrically regular deposit may be suitable for

mining using a mechanized, mass-mining method, such as block caving. A small deposit of

the same ore type may require selective mining and precise ground control to establish a

profitable operation. In addition to its direct significance, there is also an interrelation

between ore body size and the other geometric properties of configuration and disposition, in

their effect on mining method.

6. Geomechanical setting

The geo-mechanical setting includes:

Rock material properties such as strength, deformation characteristics (such as

elastic, plastic and creep properties) and weathering characteristics.

Rock mass properties are defined by the existence, and geometric and mechanical

properties, of joint sets, faults, shear zones and other penetrative discontinuities.

The pre-mining state of stress in the host rock is also a significant parameter.

In addition to the conventional geomechanical variables, a number of other rock material

properties may influence the mining performance of a rock mass. Adverse chemical

properties of an ore may preclude caving methods of mining, which generally require

chemical inertness. For example, a tendency to re-cement, by some chemical action, can

reduce ore mobility and promote bridging in a caving mass. Similarly, since air permeates a

caving medium, a sulphide ore subject to rapid oxidation may create difficult ventilation

conditions in working areas, in addition to being subject itself to degradation in mechanical

properties.

Other more subtle ore properties to be noted are the abrasive and comminutive properties of

the material. These determine the drillability of the rock for stoping purposes, and its particle

size degradation during caving, due to autogeneous grinding processes. A high potential for

self-comminution, with the generation of excessive fines, may influence the design of the

height of draw in a caving operation and the layout and design of transport and handling

facilities in a stoping operation.

In some cases, a particular structural geological feature or rock mass property may impose a

critical mode of response to mining, and therefore have a singular influence on the

appropriate mining method. For example, major continuous faults, transgressing an orebody

and expressed on the ground surface, may dictate the application of a specific method, layout

and mining sequence. Similar considerations apply to the existence of aquifers in the zone of

61

potential influence of mining, or shattered zones and major fractures which may provide

hydraulic connections to water sources. The local tectonic setting, particularly the level of

natural or induced seismic activity, is important. In this case, those methods of working

which rely at any stage on a large, unfilled void would be untenable, due to the possibility of

local instability around open stopes induced by a seismic event. A particular consequential

risk under these conditions is air blast, which may be generated by falling stope wall rock.

7. Orebody value and spatial distribution of value

The monetary value of an orebody, and the variation of mineral grade through the volume of

the orebody, determines both mining strategy and operating practice. The critical parameters

are average grade, given various cut-off grades, and grade distribution. The average grade

determines the size and monetary value of the deposit, since the market price for the mineral

changes with time and demand.

The significance of dilutions of the ore stream, arising, for example, from local failure of

stope wall rock and its incorporation in the extracted ore, is related to the value per unit

weight of ore. In particular, some mining methods are prone to dilution, and marginal ore

may become uneconomic if mined by these methods. Grade distribution in an orebody may

be uniform, uniformly varying (where a spatial trend in grade is observed), or irregular

(characterized by high local concentrations of minerals, in lenses, veins or nuggets). The

concern here is with the applicability of mass mining methods, such as caving or sublevel

stoping, or the need for complete and highly selective recovery of high-grade domains within

a mineralized zone. Where grade varies in some regular way in an orebody, the obvious

requirement is to devise a mining strategy which assures recovery of higher-grade domains,

and yet allows flexible exploitation of the lower-grade domains.

Engineering environment

8. Engineering Environment

A mining operation must be designed to be compatible with the external domain and to

maintain acceptable conditions in the internal mining domain. Mine interaction with the

external environment involves effects on:

Local groundwater flow patterns, changes in the chemical composition of

groundwater,

Possible changes in surface topography through subsidence. In general, caving

methods of mining have a more pronounced impact on subsidence than supported

methods.

Mine gases such as methane, hydrogen sulphide, sulphur-dioxide, carbon dioxide or

radon may occur naturally in a rock mass, or be generated from the rock mass during

mining activity.

62

In fact, stope backfill generated from mill tailings is an essential component in many mining

operations. Specific mining methods and operating strategies are required to accommodate

the factors which influence the mine internal environment.

Problems

Q1. Discuss the effects of rock mass response to stoping?

Q2.Explain how rock mass movement due to stoping affect ore dilution in different

stoping operations?

Answer:

Dilution is defined as the low grade (waste or backfill) material which comes into an ore

stream, reducing its value. By-and-large, dilution control may be more difficult in the caving

methods where displacements of large magnitudes within the host rock are experienced.

Artificially supported mining methods rely on achieving close control of the performance of

the rock mass surrounding a stope. Cut and fill relies on passive support from the applied

backfill, while shrink and VCR stoping use the broken ore as a temporary support for the

stope walls. Shrinkage stopes can be susceptible to external dilution due to time dependent

failure of the exposed walls, while excessive damage (external dilution) to the stope walls can

be experienced during VCR mining, specially when used for pillar recovery.

The success of naturally supporting methods such as sublevel open stoping (for large tabular

and massive ore-bodies) relies on achieving large stable and mostly unsupported stope

boundaries. The stand-up time before backfill support is introduced as well as support

provided by cable bolting is also an important factor controlling stability.

(Source of information: Ernesto Villaescusa)

Q3.What technical information is needed for preliminary mine planning?

Answer:

Many details must go into the planning of underground mine and information must come

from several sources. Geological, structural, and mineralogical information must first be

collected and combined with data on resources and reserves. This information leads to the

preliminary selection of a potential mining method and sizing mine production.

The following information should be gathered during the exploration phase and passed on to

the mine evaluation team of the mine development team. The information is:

Property location and access

Description of surface features

Description of regional, local, and mineral deposit geology

Review of exploration activities

Tabulation of potential ore reserves and resources

63

Explanation of ore-reserve calculation method

Description of company’s land position

Description of the company’s water position

Ownership and royalty conditions

History of the property

Any special studies by the exploration team

Any social issues or environmental issues that have surfaced while exploration was

being completed.

Q4. What specific planning is required related to physical properties of the ore body

and surrounding ground?

Answer:

The physical nature of the extracted rock mass and the rock mass left behind are very

important in planning many of the characteristics of the operating mine. Four aspects of any

mining system are particularly sensitive to rock properties.

(a). the competency of the rock mass in relation to the in situ stress existing in the

rock determines open dimensions of unsupported roof unless specified by

regulations. It also determines whether additional support is needed.

(b). When small openings are required, they have a great effect on productivity,

especially in harder materials for which drill and blast cycles must be used.

(c). The hardness toughness and abrasiveness of the material determines the type

and class of equipment that can extract the material efficiently.

(d). If the mineral contains or has entrapped toxic or explosive gases, the mining

operation will be controlled by special provisions in mine regulations.

64

Chapter 5 Mining Methods

The emphasis is confined to the relations between working method, the rock mass conditions

essential to sustain the method, and the key orebody properties defining the scope for

application of the method. The mining methods commonly employed in industrial practice

are classified as shown below. Other mining methods, mostly of historical or local

significance, such as top slicing or cascade stoping, could be readily incorporated in this

categorization. The gradation of rock performance, ranging from complete support to induced

failure and granular flow, and in spatial energy change from near-field storage to far-field

dissipation, is consistent with the notions discussed earlier.

Classification of stoping methods based on the strength of the rock mass

A. Naturally supported stopes

1. Open stoping with pillar supports

a. Room-and-pillar stopes

Room-and-pillar with regular pillars

Room-and-pillar with irregular pillars

2. Open stopes

a. Sub-level open stoping

b. Large Diameter Blast Hole stoping (Long hole stoping)

B. Artificially supported stopes

3. Shrinkage stoping

a. With pillar (post pillar)

b. Without pillars

c. With subsequent back filling

4. Cut-and-fill stoping

a. Horizontal cut-and-fill stoping

b. Post pillar cut-and fill stoping

5. Vertical Crater Retreat – with back filling

6. Square set stoping

C. Caved stopes

7. Sub-level caving

8. Block caving

A summary of factors for each U/G mining method, including the suitable orebody

geometries, orebody grades, orebody and country rock strengths, and depths are shown in

Table 1.

65

Table 1: Summary of geotechnical factors for each underground mining method

Method

Class Method

Relative

magnitude of

displacements

in country

rock

Strain

energy

storage

in near

field rock

Suitable

orebody

geometry

Suitable

orebody

grade

Suitable

orebody,

country rock

strength

Suitable

depth

Pillar

supported Room-and-pillar Very low

Very high

Tabular,

maximum

dip 55°

High

Both strong

and

competent,

low frequency

of cross

jointing in

roof

Shallow

Pillar

supported

Sublevel open

stoping Very low Very high

Massive or

steeply

dipping

stratiform,

regular

boundary

Moderate

Must be

sufficient to

provide stable

walls, faces,

and crown for

stopes

Variable

Artificially

supported Cut-and-fill Low High

Veins,

inclined

tabular,

massive;

35-90° dip

High;

variable

with lenses

is

acceptable

Competent

orebody, can

be weaker

country rock

Shallow

or deep

Artificially

supported Bench-and-fill Low High

Narrow

vein

mining

High

Competent

orebody, can

be weaker

country rock

Shallow

or deep

Artificially

supported Shrink stoping Moderate Moderate

Narrow

extraction

blocks;

veins,

inclined;

tabular,

massive

High;

variable

with lenses

is

acceptable

Competent

orebody (and

resistant to

crushing), can

be weaker

country rock

Shallow

or deep

66

Artificially

supported VCR stoping Moderate Moderate

Minimum

3 m width

orebody;

veins,

inclined

tabular,

massive

High;

variable

with lenses

is

acceptable

Competent

orebody (and

resistant to

crushing), can

be weaker

country rock

Shallow

or deep

Unsupported Sublevel caving High Low

Steeply

dipping ore

bodies

High

enough to

sustain

dilution

(perhaps

>20%)

Reasonably

strong

orebody rock

enclosed by

weaker

overlying and

wall rocks

From

shallow

to deep

Unsupported Block caving Very high Very low Large ore

bodies

where

height

>100 m

High

enough to

sustain

dilution

Rock mass of

limited

strength, with

at least two

prominent

sub-vertical

and one sub-

horizontal

joint set

Shallow

or deep

1. Naturally Supported Method- Room-and-Pillar Mining

A mining method based on natural support seeks to control the rock mass displacements

through the zone of influence of mining, while mining proceeds. This implies maintenance of

the local stability of the rock around individual excavations and more general control of

displacements in the near-field domain. (Ref: Brady & Brown1993).

Conditions

• Ore strength: weak to moderate

• Host rock strength: moderate to strong

• Deposit shape: massive; tabular

• Deposit dip: low (< 35 degrees), preferably flat

• Deposit size: large extent – not thick

• Ore grade: moderate

67

Features

• Generally low recovery of resource as pillars needs to be left (40-60%)

• Moderately high production rate

• Recovery can be improved with pillar extraction (60-80%) but caving and

subsidence will occur

• Suitable for total mechanization, not labour intensive

• High capital cost associated with mechanization

• Versatile for variety of roof conditions

Applications

• Room and pillar mining – eg. Agnigundala Lead-Zinc mine of HZL,

Tummallapalli Uranium Mines of UCIL

• Variation: Stope and pillar mining

Stope development;

In-stope raises – minimum two as per the regulation, so that one raise acts as a

ventilation in-take raise and the other the return. (eg.2x2 m raise dimension)

The level interval decides the width of the stope - that is the length between

the upper and lower level.(eg. 30 – 60 m level interval)

The length of the stope, i.e the distance between the terminal raises of a stope;

it is also known as the block size and it is usually as per the grade value of the

ore deposit. (eg. 60m – 100m)

Ore draw point development. – Ore drawing is based on the degree of

mechanization of the mine. Eg. The ore-drawl in UCIL mines is by LHD (load

Haul Dumpers) and LPTD (Low Profile dump Trucks). The LPDTs move into

the stope and carry the material through a ramp to the main ore pass.

Fig . Low Profile Dump Truck (LPDT)

Method:

The room and pillar mining method is a type of open stoping used in near horizontal

deposits in reasonably competent rock, where the roof is supported primarily by pillars. Ore

is extracted from rectangular shaped rooms or entries in the ore body, leaving parts of the ore

between the entries as pillars to support the hanging wall or roof. The pillars are arranged in a

68

regular pattern, or grid, to simplify planning and operation. They can be any shape but are

usually square or rectangular. The dimensions of the rooms and pillars depend on many

design factors. These include the stability of the hanging wall and the strength of the ore in

the pillars, the thickness of the deposit, and the depth of mining. The objective of design is to

extract the maximum amount of ore that is compatible with safe working conditions. The ore

left in the pillars is usually regarded as irrecoverable or recoverable only with backfill. In this

case backfill costs or the potential loss of valuable resource may be a limiting factor in room

and pillar mining at greater depths.

The principal advantage of room and pillar stoping is that it is readily adaptable to

mechanized mining equipment, which results in high productivity and a relatively low cost

per ton of material extracted. For large ore bodies, a large number of working places can be

easily developed so that high daily rates of production can be counted upon. Most of the mine

development work is in ore, so waste extraction is kept to a minimum.

Figure Elements of a Room-and-Pillar stoping method

69

Figure Ore handling in a Room-and-Pillar stoping method.

The main disadvantage of room and pillar mining is that a large area of roof is continuously

exposed where work activities or movement of men and supplies are carried out.

Consequently, roof condition is a primary concern for the safety of personnel and ground

support is generally a major cost, especially in rooms with high backs. Also, recirculation of

ventilating air can be difficult to minimize in room and pillar mines.

Components of a supported mine structure

A mining method based on pillar support is intended to control rock mass displacements

throughout the zone of influence of mining, while mining proceeds. This implies maintenance

of the local stability of rock around individual excavations and more general control of

displacements in the mine near-field domain. As a first approximation, stope local stability

and near-field ground control might be considered as separate design issues. Near-field

ground control is achieved by the development of load-bearing elements, or pillars, between

the production excavations. Effective performance of a pillar support system can be expected

to be related to both the dimensions of the individual pillars and their geometric location in

the orebody. These factors are related intuitively to the load capacity of pillars and the

loads imposed on them by the interacting rock mass.

Room-and-Pillar stoping method

70

Fig Plan view of a Room-and-Pillar stope

Fig. Samsung limestone mines – South Korea

71

Analysis of Pillar support system in Room-and –Pillar stoping

A mining method based on pillar support is intended to control rock mass displacements

throughout the zone of influence of mining, while mining proceeds. This implies maintenance

of the local stability of rock around individual excavations and more general control of

displacements in the mine near-field domain. Near-field ground control is achieved by the

development of load-bearing elements such as pillars, between the production excavations.

Effective performance of a pillar support system is related to:

1). the properties of the material,

2). geological structures,

3). absolute and relative dimension of the pillar,

4). the nature of surface constraints applied by the country rock,

5). geometric location of the pillars in the orebody.

These factors are related to the load capacity of pillars and the loads imposed on them by the

interacting rock mass. Since a lot of ore remains locked-up in the pillars, an economic design

Fig Plan view of a Room-and-Pillar stoping method

Fig Plan view of a Room-and-Pillar stoping method

72

suggests that ore locked-up in pillars be a minimum, while fulfilling the essential

requirement of assuring the global stability of the mine structure. Therefore, detailed

understanding of the properties and performance of pillars and pillar systems is

essential in mining practice, to achieve the maximum, safe economic potential of an orebody.

Figure. Schematic illustration of problems of mine near-field stability and stope local

stability, affected by different aspects of mine design.

In a classical Room-and-Pillar stoping method, pillars in flat-lying, stratiform ore-bodies are

frequently isolated on four sides, providing a uniaxial loading condition from the

hang-wall rock mass. Interaction between the pillar ends and the country rock results in

heterogeneous, triaxial states of stress in the body of the pillar, even though it is uniaxially

loaded by the abutting rock.

The figure below illustrates the types of pillars in an ideal room-and-pillar stope.

Fig. Layout of barrier pillars in a room-and-pillar stope(Ref. Rock Mech. for u/g mining Brady & Brown)

In order to restrict the stope instability limited to a single room-and-pillar stope, the adjacent

stopes are separated by a barrier pillar, similar to the division of panels in a coal mine. The

barrier pillars are designed such that each stope (panel) performs as an isolated mining

73

domain. The maximum extent of any collapse is then restricted to that stope pillars itself. The

stope stability is therefore controlled by the response of stope pillars in a room-and-pillar

stope. A set of uniaxially loaded pillars is illustrated in the Figure below.

Fig. Room-and-Pillar stope, pillar configuration

Fig. Room-and-Pillar layout showing load carried by a single pillar assuming total load to

be uniformly distributed over all pillars(Ref. Hoek & Brown Underground excavation in rock)

Fig Average pillar stress in room-and-pillar stope

74

Figure Redistribution of stress in the axial direction of a pillar.

Modes of Pillar failure

Stoping activity in an orebody causes stress redistribution and an increase in pillar loading,

illustrated conceptually in Figure above. For states of stress in a pillar less than the in situ

rock mass strength, the pillar remains intact and responds elastically to the increased state of

stress. Mining interest is usually concentrated on the peak load-bearing capacity of a pillar.

Subsequent interest may then focus on the post-peak, or ultimate load-displacement

behaviour, of the pillar. The structural response of a pillar to mining-induced load is

determined by the rock material properties, the geological structure, the absolute and relative

dimensions of the pillar and the nature of surface constraints applied to the pillar by the

country rock. Three main modes of pillar behaviour under stresses approaching the rock mass

strength have been recognized, which may be reproduced qualitatively by laboratory tests on

model pillars.

Different modes of failure as seen in the ffield observations are:

1. Fretting or or necking of the pillar: Fretting occurs in relatively massive rock with

moderately strong H/W, F/W, and ore body. One of the main causes for necking is the

development of tri-axial stress condition at the wall contacts (H/W and F/W), which

result in the development of shear stresses at the contact zones and the failure is

localised in the central part of the pillar. The failure is due to tensile stress

concentration. The most obvious sign of pillar stressing involves spalling from the

pillar surfaces, which consequently leads into the development of hour-glass shaped

pillar.

Fig. Fretting (Samsung Limestone Mines – South Korea)

Original pillar

boundary

spalling

75

2. Shear failure: The effect of pillar relative dimensions on failure mode is illustrated in

the second most common failure- which is shear failure along a shear plane. For

regularly jointed orebody rock, a high pillar height/width ratio may favour the

formation of inclined shear fractures dividing the pillar across plane of weakness.

There are clearly kinematic factors promoting the development of penetrative,

localized shear zones of this type. Their occurrence has been reproduced in model

tests by Brown (1970-Ref. Rock Mech. Brady & Brown ), under the geometric

conditions prescribed above.

Fig. Failure along a shear plane (Samsung Limestone Mines – South Korea)

3. Axial Splitting (Bulging or barrelling): The third major mode of pillar response is

seen in an ore body which is relatively strong in comparison to the wall rocks and

hang-wall rocks form highly deformable plane of weakness at the contact plane of the

pillars. The relative deformation of the pillar and the hang-wall rocks generates

transverse tractions over the pillar end surfaces and promotes internal axial splitting of

the pillar. This may be observed physically as lateral bulging or barrelling of the pillar

surfaces. Geomechanical conditions favoring this mode of response may occur in

stratiform orebody, where soft bedding plane partings define the foot wall and

hanging wall for the ore-body. The failure condition is illustrated in Figure 13.5c.

Fig. Splitting of pillars ( Barrelling/ bulging)

4. Structural failure: This mode of failure is commonly seen in layered ore bodies,

such as limestone or banded hematite quartzite (BHQ). The response of the failure to

the super incumbent load is related directly to the structural geological features of the

pillar. A pillar with a set of natural fractures or bedding planes forms the weak planes

for the fracture initiation along these planes of weakness. The failure is similar to the

shear failure, where in slip takes place when the shearing stress on these planes is

more than the frictional resistance.

Shear Plane Shear plane

76

Fig. Structural failure of the pillar.

5. Buckling of Pillars: This is common in slender pillars, where width/height ratio of

the pillars is very less (0.4 -0.5).A slender pillar with well-developed foliation or

schistosity parallel to the principal axis of loading will fail in buckling mode, as

shown in the figure below.

Figure Buckling mode of deformation of pillars

Figure. Mode of fracture and failure in mine pillar

Bedding planes normal to the

loading axis

77

Measures to control the Pillar failure

Table Rock mass classification of Pillars in limestone mines (ref. Pillar stability issues based on a survey of

pillar performance in underground limestone mines; 25th international conference on ground control in mines, Gabriel S. Esterhuizen etal)

Some of the common methods of preventing the pillar failure in room-and-pillar stoping are:

1. Back filling the stope, the fill material surrounding a pillar may act as a confining

material and hence prevents the failure of the pillars.

Figure Plan view of a room-and-pillar stope

2. Rock bolting or lacing the pillar.

78

Partially benched pillar failing

under elevated stresses at the

edge of bench mining. Typical

hourglass formation indicates

overloaded pillar. The width-

to-height ratio is 0.44 based on

full benching height and the

average pillar stress is about

12% of the UCS. (Ref. Pillar

and Roof Span Design

Guidelines for Underground

Stone Mines Gabriel S.

Esterhuizen, Dennis R.

Dolinar, John L. Ellenberger,

and Leonard J. Prosser ; IC

9526; NIOSH;2011)

Partially benched pillar

that failed along two

angular discontinuities.

Width-to-height ratio is

0.58 based on full

benching height; average

pillar stress is about 4% of

the UCS.

Pillar that had an original

width-to-height ratio of 1.7,

but failed by progressive

spalling. Thin, weak beds

are thought to have

contributed to the failure.

The average pillar stress

was about 11% of the UCS

prior to failure.

Stable pillars in a limestone mine at a depth of cover of 275 m (900 ft). Slightly concave pillar ribs formed as a result of minor spalling of the hard,

Pillar that has been clad with chain link mesh to prevent further deterioration of the ribs.

79

Pillar stress estimation by tributary area method

The term tributary area method is used for estimating the average state of axial stress in the

pillar. The area extraction ratio, R, defined as the ratio of area mined to total area of ore body.

Considering the representative element of the ore body illustrated in the figure above, the

area extraction ration is also defined by

Figure below shows a cross section through a flat-lying orebody, of uniform thickness,

being mined using long rooms and rib pillars. Room spans and pillar spans are

Wo and Wp respectively.

Figure. Tributary area method to calculate the average pillar stress (ref. Brady & Brown)

Considering the requirement for equilibrium of any component of the structure under the

internal forces and unit thickness in the anti-plane direction, the free body shown in the figure

below yields the following equation

On considering equilibrium,

𝜎𝑃𝑊𝑃 = 𝑃𝑧𝑧(𝑊𝑜 + 𝑊𝑃)

Or

𝜎𝑝 = 𝑃𝑧𝑧(𝑊𝑜 + 𝑊𝑝)/𝑊𝑝

In this expression, is the average axial pillar stress and is the vertical normal component of

the pre-mining stress field. The width (of the representative free body of the pillar structure is

often described as the area which is tributary to the representative pillar. The term tributary

area method is therefore used to describe this procedure for estimating the average state of

axial stress in the pillar. The area extraction ratio, r, defined as the ratio of area mined to total

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area of ore body. Considering the representative element of the ore body illustrated in the

figure above, the area extraction ration is also defined by

𝑟 = 𝑊𝑜/(𝑊𝑜 + 𝑊𝑃)

So that

1 − 𝑟 =𝑊𝑃

𝑊𝑜 + 𝑊𝑃

Insertion of this expression in the above equation, yields:

𝜎𝑃 = 𝑃𝑧𝑧[1

1 − 𝑟]

The mining layout shown in the following figure, involving pillars of plan dimensions a and

b, and rooms of span c, may be treated in an analogous way.

The area tributary to a representative pillar is of plan dimensions (a+c), (b+c), so that

satisfaction of the equation for static equilibrium in the vertical direction requires

𝜎𝑃𝑎𝑏 = 𝑃𝑧𝑧(𝑎 + 𝑐)(𝑏 + 𝑐)

Or

𝜎𝑃 =𝑃𝑧𝑧(𝑎 + 𝑐)(𝑏 + 𝑐)

𝑎𝑏

The area extraction ratio is defined by

𝑟 =[(𝑎 + 𝑐)(𝑏 + 𝑐) − 𝑎𝑏]

(𝑎 + 𝑐)(𝑏 + 𝑐)

With some simple rearrangement the above equation yields the following:

𝜎𝑃 = 𝑃𝑧𝑧[1

1 − 𝑟]

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For a square pillar, of plan dimension WpxWp, are separated by rooms of dimension Wo, the

equation is

𝜎𝑃 = 𝑃𝑧𝑧[(𝑊𝑜 + 𝑊𝑃)/𝑊𝑃]2

The pillar stress expression given above helps in a rough estimation of the pillar stresses.

Fig. Relationship between the pillar stresses and the area extraction ratio

The relationship between the pillar stress and the area extraction ratio is illustrated in the

above figure. The main observations from the above graph are that:

1. The average pillar stress is directly proportional to the area extraction ratio and the

relationship is non-linear.

2. There are two distinct zones in the above relationship, where in the increment in the

pillar stress until r = 0.75, is near linear and the slope is mild, whereas the nonlinear

exponential increment is seen beyond a point where r > 0.75.

3. In the second zone, a very small increase in the extraction ratio is developing a high

increment in the pillar stress.

4. It is therefore inferred that for keeping the stope stable, it is imperative that the

extraction ratio needs to be within the limits of tolerable stress concentration levels, in

the pillar (Factor of safety of the pillar is > 1).

Limitations of Tributary area method

1. The stress estimated by this method represents an average stress within the pillar, and

it is purely a convenient way of representing the state of loading of a pillar in a

direction parallel to the principal stress.

2. Tributary area analysis restricts attention to the pre-mining normal stress (in-situ

stress) component directed parallel to the main axis of the pillar support system.

3. It is assumed that the effect of other stresses in other direction have no effect, which

in reality is not always true.

4. The stress coming on the pillar is the induced stress.

5. Strength of the pillar is related to its volume and geometric shape.

r= 0.75r= 0.9

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6. Tributary area method provides a simple method of determining the average state of

axial stress in a pillar. The size of the pillars is bigger in the mines, say 4x4 or 4x6, or

6x8 and so on. Increasing the volume of the pillar increases the number

discontinuities but the shape of the pillar may give rise to the effect of confinement to

the core pillar.

Fig. Distribution of vertical stresses in a pillar (Ref. – Brady & Brown Rock Mech- Wagner-1980)

The measurement of the load distribution in a pillar at various states of loading is shown in

the above figure. It is seen from the above figure that the failure commences from the

boundaries of a pillar and migrates towards the centre (core pillar). It may so happen that the

structural failure of the pillar has occurred but the core pillar has not reached its full load-

bearing potential.

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2. Sublevel Stoping

Sub-level Stoping is an open stoping method applicable to relatively medium-wide ore bodies

of steep dip where broken ore easily moves down to the draw points by gravity. Both ore and

wall rocks need to be strong as the stope has to stand until the broken ore has been drawn out

of the stope and the rib and crown pillars are also recovered eventually.

This method requires more stope development work compared to some other methods like

shrinkage stoping or room-and-pillar mining (15% of ore comes from development workings

in sub-level stoping as compared to 10% in other methods), but this does not affect the cost of

mining as most development is in ore. Indeed sub-level stoping is a low cost mining method

comparable to caving methods. With the development of long hole drilling practice sub-level

interval has progressively increased thus bringing down the stope development work. On the

other hand long hole blasting has restricted the method to fairly regular ore bodies so as to

minimize dilution.

In this method the orebody is vertically divided into levels, and between two levels the stopes

of convenient size are formed. A rib pillar is left in between them separates two adjacent

stopes. Leaving a crown pillar at the top of the stope protects the level above, whereas lower

level is used as haulage level to gather the ore from the stopes. Vertically the stope is divided

into a number of horizons by suitably positioned drill drives called the sublevels, and hence

the name sublevel stoping. With advent of new drill machines with the ease of drilling large

diameter blast holes, the conventional sub-level stoping method has gone through lot of

modifications. When the drills used for the purpose of stope drilling are the blast-hole drills,

as such, sometimes this is also known as “blast-hole” stoping method with large diameter

blast holes, it is named as Large Diameter Blast Hole Stoping – LDBH method. Based on the

thickness of the orebody and the orientation of the levels the Sub-level stoping is known as:

Longitudinal sub-level stoping method,

Transverse sub-level stoping method

Conditions suitable for Sub-level stoping

1. Geotechnical parameters:

Ore strength: moderate to strong (> 40 MPa UCS)

Host rock (Footwall and hang wall rocks) are also strong

2. Geometry, disposition & orientation:

Deposit shape: tabular or lenticular, regular dip and defined boundaries

Deposit dip: steep (>45-50 degrees, preferably 60-90 degrees)

Deposit size: 6-30m wide, fairly large extent

(Very thin deposits : 0.7m; Thin deposits; 0.7 – 2m;

Medium thick: 2m- 5m; Thick deposits: 5 – 20 m; Very thick > 20 m;

eg. HCL- Malanjkhund copper deposit is 80 m thick )

Ore grade: moderate

84

Figure Transverse and longitudinal sections of sub-level stoping method

Stope Development:

Stope development varies with the method of ore drawing from the stope, stope width and

inclination. There are four methods of ore drawing from the stope to the gathering or haulage

level. Ore broken in the stope is collectd in a series of mill holes which are conical in shape

with a side slope of 450. However, it is more time consuming to prepare individual mill holes

so that it is a common practice to-day to replace a row of mill holes by a continuous trough

which is excavated from trough drive by parallel fans of upward long holes drilled from the

trough drive.

Ore from mill holes or trough passes down to draw point cross-cuts. Secondary blasting of

boulders, if required is done in the draw point cross-cut. The draw point cross-cuts can

directly lead to the haulage level where ore loading is done by loaders into mine cars. But

with the introduction of LHDs it is a common practiced today to connect the draw point

cross-cuts to a gathering drive which is in turn connected to the haulage level below through

a transfer raise.

Draw point loading with high capacity LHDs (2.5 – 3.8 bucket capacity) is the best choice for

high production. Even here production can be maintained only with good fragmentation in

primary blasting so that disruption of loading at the draw point due to boulder blasting is

85

minimized. However, where availability of LHDs is poor a gravity transfer system with a

gizzley level is preferable because of its reduced cost of ore transfer.

Sub-levels are driven between raises at the ends of the stope block. A single sublevel at a

horizon suffices upto 15m width of ore body but beyond this two sulevels are usually driven

from a cross cut from the raise. In not too steep ore bodies, a sublevel placed at the footwall

helps in breaking a smooth footwall for easy filling down of the ore. Sublevel interval

generally varies from 10 to 20 m with haulage levels placed 50 – 90 m apart.

While larger sublevel intervals have the advantages of less sublevel drivage per unit

production and less unit drilling time or faster overall drilling rate resulting from longer

holes being drilled from a single setting of the drill, they have the following disadvantages:

rate of penetration slows down with longer holes due to energy attenuation at a larger

number of steel junctions (with the commonly used drifters and prevalent air pressure

a maximum hole length of 30 m is desirable)

hole deviation may become significant

with a larger sublevel interval the number of holes in a ring increase in order to

maintain the required toe burden which results in crowding of the holds at the cellar.

open space required at the undercut level to accommodate the swell of blasted ore

increases with increased sublevel spacing.

Length of stope ranges from as small as 20m to as large as 150m depending on

ground pressure and rock characteristics. Rib pillars vary from 6 to 10 m length.

Figure Sub-Level Stoping method

86

Figure Ore drawal in sub-level stoping

Stoping:

Stoping starts from a slot at one end of the stope sometimes a slot is made at the centre of the

stope, but this is desirable in relatively long stopes which is possible in relatively strong

ground. The slot is made by stripping a slot raise from wall to wall by parallel long holes drill

from cross-cuts it the ends of the upper sublevels. Rings of long holes are then drilled from

the sublevels and blasted into the slot.

Rings of 45 to 65 holes (52 & 57 mm holes are most prevalent) are usually drilled. Burden on

rings is generally kept at 30 times hole diameter while a larger toe spacing 50 times hole

diameter is adopted between holes in a ring. This should however, be varied depending on

rock structure and strength and fragmentation desired based on experimental stope blasts.

Blasting efficiency requires a correct maintenance of the drilling pattern through rigorous

control of hole direction and deviation.

Charging of holes also needs careful planning to ensure efficiency in blasting and good

fragmentation. Holes should be charged to different distances from the collar in order to

maintain a uniform charge spacing as far as practicable. Charging pattern must be worked out

in the planning office and supplied to the face.

87

Parallel hole drilling is superior to ring drilling. It ensures lower specific drilling and blasting

compared to ring drilling. However, for safety it is desirable to drill the parallel holes from a

cross-cut rather than from a bend. With narrow holes this loads to a large burden on holes and

hence poor fragmentation in addition to the need for driving narrow cross-cuts at very close

intervals. Use of large diameter blast holes however, has made parallel hole drilling a

superior proposition. This will be dealt with further under blast-hole stoping.

3. Shrinkage Stoping

Ground Conditions

Ore strength: strong (other characteristics important – should not pack, oxidise

Or spontaneously combust)

Host rock strength: strong to fairly strong

Deposit shape: tabular or lenticular, defined boundaries

Deposit dip: steep(>50 degrees or angle of repose)

Deposit size: 1-30m wide – fairly large extent

Ore grade: fairly high

Features

Suited to smaller scale operations –moderately low production

Labour intensive, dangerous work conditions

Low capital investment

Moderately selective

Majority of ore tied up in the stope

Ore subject to oxidation, packing and spontaneous combustion in stope

Variations: Vertical Crater Retreat

Fig. Longitudinal and cross section of shrinkage stoping

88

Fig. Shrinkage stope layout for Load-Haul-Dump (LHD) operations

89

Fig. Shrinkage stope layout for slusher trench operations

4. Cut-and-Fill Stoping

Supported class of methods consists of those methods which require substantial amount of

artificial support to maintain stability in exploitation openings and systematic ground control.

The one of the supported class in common use today is cut and fill method.

In this method the ore is excavated in horizontal slices starting from the bottom of the stope

and advancing upwards. The broken ore is loaded and completely removed from the stope.

When ore slice of the ore has been excavated the corresponding volume is filled with waste

material. The filling is conducted integrally with the mining cycle and not after the

completion of the entire mining operation.

The filling material can consist of waste rock from preparation, distributed mechanically over

the stoped out area. In modern cut and fill method however the hydraulic filling method is

normal practice. The filling material here consists of fine grained tailing from ore dressing

plant (mill tailing) or sand mixed with water transported into the mine and distributed through

pipe lines. When the water is drained off a competent fill with a smooth surface is produced.

90

Special drainage technique is required since the slivery contains 30-40% water. The tops of

man ways and ore passes must be extended above the fill floor to keep them open. To provide

proper drainage of the fill while it sets percolation drains (perforated pipes) are installed

along the stope sill and decantation towers are maintained through the fill; run-off-water must

be disposed off in the drainage system on the level below:

Figure Longitudinal section of a Cut-and-Fill Stope

Fig Cut-and-Fill stoping operation

91

Fig Cut-and-Fill stoping operation

Features

Low development cost

High mining cost, due to backfilling operations

Permits good selectivity, is versatile, flexible and adaptable

Backfilling can disrupt mining operation

Labour intensive

Application :

This method can be used with steeply dipping ore bodies with reasonably firm ore. The

condition can be stated as:

Ore strength: moderate to strong, maybe less competent than with un-

supported method.

Rock Strength: Weak

Deposit Shape: tabular, can be irregular

Dip: moderate to fairly steep can accommodate flatter deposit if ore passes

are steeper than angle of repose

Deposit Size: narrow to moderate width

Ore grade: fairly high

Depth: moderate to deep

Advantages :

1. Moderate productivity

2. Moderate production rate

3. Permits good selectivity, sorting possible

4. Low development cost

5. Adoptable to mechanization

6. Recovery is high, low dilution

7. Waste revised as fill

92

8. Good safety record.

Disadvantage :

1. Fairly high mining cost

2. Handling of waste – extra cost

3. Filling complicates cycle of operation causing discontinuous production

4. Must provide stope access for mechanized equipment

5. Compressibility of fill risks some ground settlement

Preparation :

Preparation consists of :-

Haulage drift along the stope

Undercut of the stope usually 5 – 10m above the haulage drift

Short raises for man-ways and ore passes from haulage drift to undercut

Raise from undercut to level above for fill transport and ventilation.

Cycle of operation:

Drilling:

The ore slice can be drilled in two different ways, with horizontal stope holes or with upward

holes with the later method on certain headroom is required between the back and the fill

surface usually 2-2.5m. After blasting and removal of the ore the distance is increased to 5 -

6m which means that a competent ore and hanging wall is needed.

For the drilling light rock drills are often used though mechanized jumbos are also used. An

advantage of upward drilling method is that large sections of the roof can be drilled without

interruptions and large round can be blasted.

Fig Cut-and-Fill stoping operation

93

Blasting:

Blasting round consists of horizontal or inclined holes and charging the cartridge or slurries.

Loading :

Previously scraping was used in the stope to bring the blasted muck from two ends upto the

central ore pass. Since able to work directly on a hydraulically filled surface auto loaders are

very suitable for loading in stopes where the operation is characterized by a short haul. In

comparison with scrapers these loaders are more versatile, clean the stope efficiently and

work unaffected by curves and supports.

In highly mechanized stopes with hydra boom jumbos for drilling, the loading and transport

are often done with heavy diesel front end loaders or LHD’s. The average distance of travel is

generally 60m but has also been as high as 240m. This has made it possible to space the ore

passes far apart and save their cost.

Post Pillar Stoping:

Post Pillar Cut-and-Fill stoping is adopted in such cases where the width and/or inclination of

the ore-body or the condition of hanging wall, stope back or induced stresses are such that

ordinary methods of rock bolting and fill would not give sufficient support to the stope back.

The rock mechanics aspect of the mining method is that post pillars are expected to yield

under the fill and be constant in the post-failure condition, most of the super-incumbent load

being transferred to the abutments. The pillars, however, surrounded by the backfill must still

have sufficient strength to support the immediate roof structure. The role of the fill in this

situation is to prevent any further disintegration or unravelling so that pillars can continue to

support the immediate stope back strata rather than the total overlying strata.

This method of stoping was adopted in Mosabani copper mines of HCL. In this mine the

post-pillar stoping method, 4 x 4 m square narrow post pillars were left at regular intervals

with clear room of 13 m along the strike and 9 m across it. The desing was based on the

assumption that the pillars would give additional strength to the stope back and the fill

materials surrounding the narrow pillars, provide them with lateral support and maintain their

integrity. But in actual practice, at deeper levels, for example at the 27th level where post-

pillar stopes were at 950 m depth from the surface, post pillars were larger in size, mostly

varying from 4.5 x 5.5 m to 5.5 x 6.5 m and were closely spaced. The clear room in the strike

direction varied from 10 to 13 m and in the dip direction it varied from 6 to 9 m.

Approach to Post Pillar Method:

The approach was to see if the pillar support could be reduced to a point where the amount of

material being left in the pillar for support is unrecoverable. The advantages possible are:

1. the whole area can be developed simultaneously

2. future pillar mining would be eliminated resulting in maxm production rates and

minimum crew size being maintained for

3. a large portion of pillar ore would be mined using a primary method out primary

costs.

94

4. each successive elevation would be mined from footwall tu HW utilizing existing

service development

However major problems are,

design of the pillar size

determining geometry of the stope

economic problem of ore being left unrecoverable

Fig. layout of the post-pillar Cut-and-Fill stope

The problem of pillar design can be reduced to a realistic evaluation of a safety factor. The

safety factor Fs is defined as :-

Fs = CP/P

95

Where:

Fs = factor of Safety

Cp = pillar strength

P = the pillar load.

Fs > 1 →stable structure

Fs < 1 →unstable structure

However it has been proved that rock is capable of sustaining strength after failure – defined

as post failure strength. The application of post failure strength is of fundamental importance

to the successful development of the post pillar technique. For an economic extraction it was

considered that the post pillars will be loaded to failure and then gradually fail in the fill

without causing a sudden collapse. Therefore the first fundamental principle of pillar design

is “the safety factor for a post pillar should be near unity”.

Stability of Post-Pillar Cut-and-Fill stope

The strength condition of the ore body and surrounding rock imposes the application of

additional support to the back for its stability. This method like cut and fill needs support of

the immediate back and the HW by means of 1.5m long rock bolts in a systematic pattern of

1.5m x 1.5m.

Cable bolting of these stopes in geologically disturbed areas have been introduced. It is

expected to prevent sudden fall of ground, cut down the time required for placing supports

and allow production of large quantum of broken ore by taking longer cuts out a time.

All other operations are similar to that of cut and fill stopes except that unlike in cut and fill

stope the drilling is done by wagon drills and mucking is done by larger capacity LHD units.

Mechanism of pillar deformation:

A distinct feature of a post pillar is its progressive deformation with upward advance

in mining. The behaviour of the pillar is divided into two major phases: active and passive.

The active phase represents the interplay of forces due to mining (stress concentration, fill

pressure, blasting etc.). In the passive phase the deformation is complete and all the forces

have reached equilibrium. As the pillar height increases the passive phase progressively

moves upward. At the completion of mining the entire length of the pillar transforms into

passive phase.

The deformation characteristics of a pillar after six cuts may be divided into six zones.

Zone 6 is in an elastic range. Zone 5 is subjected to maximum stress concentration. The pillar

at this stage begins to fail in the exposed section (zone – 4). Zone 3 represents the post failure

response of the pillar. In zone – 2 the pillar deformation is complete. Zone 1 is primarily in

tension probably caused by yielding of a rigid pillar.

These results led to the conclusion that the stability of a pillar is only critical for the duration

of three cuts below the stope back. Its practical implication was utilized in the sequencing of

stoping blocks by maintaining a difference of 3 cuts in the adjacent stopes.

96

Fractures and spalling along the sides of some of the PPs have been observed since

their formation. As the height of these pillars increases, the condition visually deteriorates i.e.

the intensity of spalling increases and wider fractures appear. It seems that they are virtually

failed and are unable to take any load. With deterioration of the pillar condition, the stability

of stope backs in their vicinity also decreases showing sub horizontal fractures in the roof

followed by spalling.

Fig. Post-pillar cut-and-fill stope

It has been reported in the literature that at INCO’s Coleman and Falconbridge’s

Strathcona mines in Canada, an extraction of about 87% was achieved in primary stoping

leaving 6 x 6 m square post pillars and 13 x 9 m wide rooms between the pillars. The depth

below surface ranged from 550 to 700 m. Stress measurement in three post pillar stopes in

Coleman mine by the overcoring technique revealed that (i) the highest stress in the pillar

was near the abutment and (ii) the lowest stress in the pillar was near the centre of the

workings. Lateral pillar expansion measurement by extensometers showed that expansion of

the pillar was confined to the outer 1.5 m zone with no significant deformation of the central

core (3 x 3 m). Also, further 12 m mining above the level of extensometers resulted in no

additional expansion.

Rock bolting and cable bolting:

A fall of rock into excavation may occur through separation and / or shear. Separation of a

rock element from the mass of self-supporting rock is opposed by tensile resistance of the

bolts. This is achieved by anchoring the bolt in the self-supporting rock. There are cases

when there is no self-supporting rock from which anchoring can be done. Under such

97

conditions separate layers of rock can be reinforced by bolting. Thus rock bolts reinforce rock

through a fiction effect. Due to its advantage for reinforcement and easy availability rock

bolting has been the main support system both in development and stoping operations.

The bolt parameters such as type of bolt anchor, length and diameter of the bolt, pattern of

bolting and the torque to be applied on the bolts should be decided after proper examination

of rock conditions and geological disturbances. Three different types of bolts namely wedge

type, expansion type and grouted bolts are in common use.

The wedge type bolts are normally 25 mm in diameter and 1.5 m long with a slot of 15 cm on

one end and 15 cm long BSW threading on the opposite end. The dimensions of the wedge

depend on diameter of drill hole. These bolts are available in the market, but they can be

easily fabricated in any mine workshop. The wedge bolts can be easily installed by means of

jack hammer using a doly.

The expansion shell type bolts, similar to that of Pattin shell type, are also readily available in

the market. This bolt consists of 20 mm diameter and 1.5 m long mild steel rod. A hexagonal

nut is forged on one end and 20 mm BSW threading of 15 cm length is made on the other

end. A tapered wedge, having an internal thread (20 mm), can be moved over the threaded

length. Two shell halves, with serrations on the outer surfaces, are connected by means of a

bent spring steel to engage the shell halves on either side of the tapered wedge. As the rod is

rotated for tightening, the wedge comes down expanding the two shell halves and thus a grip

is provided on the rock surface for firm anchorage. As the expansion shell is to be made by

malleable casting, it will not be possible to fabricate these bolts at the mine workshop. The

expansion bolts should be installed in 37/38 mm diameter drill holes.

The advantages with the expansion bolts are (i) they can be recovered for re-use (ii) the

length of the drill hole need not be exact as in the case of wedge type bolt (the hole can be

drilled 2 or 3 cm longer than the bolt) and (iii) the bolt end will not protrude below the roof

and hence less chances of getting damaged due to blasting.

The third type of bolt available in the market is a wire mesh grouted bolt for soft strata. In

this method, the bolt is grouted through out its length whereas the other two types bolts are

anchored at the top portion of the bolt. The technique used is similar to that of Perfo bolting.

In this method, flexible wire mesh sleeve is used instead of the expensive perforated tube.

The wire mesh as well as the perforated tubes is readily available in the market for grouting

the bolts. The advantage with the grouting system is that the bolts are prevented from rusting

and corrosion and hence could be used for permanent support.

A special type of fully grouted non-tensioned bolt consists of a pair of perforated steel half

slurs of length approximately equal to that of the drill hole which are filled with a quick

setting morter wired together and inserted in the drill hole. A reinforcing bar with rounded

end is driven through the mortar filled sleeve to the bottom of the hole extruding the mortar

to fill the space out side the sleeve.

98

In the wire mesh grouted bolt similar procedure is followed. To insert the cement mortar into

the drill hole a wire mesh sleeve is used. The sleeve filled with cement mortar is inserted into

the drill hole and a ribbed steel rod is driven inside the morter. The moter came out through

the mesh and the drill hole is filled providing a firm contact to the rock surface rise of

chemical additives for the stabilization of ground support is new technology and proved a

more exact basis for roof support in difficult areas and forms a more permanent and

dependable support.

Initially epoxy resin was used but present commercial use is with polyester resin. This is

considerably less expensive, bonding is sufficient to break steel, stronger than most rocks and

cures much faster. The setting rate of polyester resin can be varied by adding an accelerator

to the mixture of resin and hardener. When put a decrease in volume of less than 2% occurs

when the resin paste becomes solid and is depended on the degree of cure. However this

volumetric shrinkage around a rock bolt is not detrimental to its effectiveness.

The commercial development of the limited injection technique has led to the fully or

partially grouted smooth bolts or deformed reinforcement bars (rebars). Full column resin

bolting has become popular with packaged resin. For full columnar grouting several

cartridges can be used with the number varied to obtain any desired length: No expansion

shell or mechanical anchor is necessary while using resin capsules because of its rapid setting

time:

The conventional rock bolt assembly consists of a smooth bolt threaded on one end with an

integral head on the other a plate and an expansion shell. Loss of anchorage at either end can

render the entire unit ineffective whereas failure of the anchorage at one point in a full

column grouting is localized and most of the length of the bolt remain effective.

Figure Different components of mechanically anchored bolts ( Brady and Brown)

99

Figure Resign bolts (Brady and Brown)

Figure : Grouted dowel in a grout –filled hole

100

Figure : Fully grouted cable bolt

Figure Summary of development of different cable bolting systems in underground

metal mines (Windsor, 1992, Brady & Brown)

101

Figure Summary of development of different cable bolting systems in underground

metal mines (Windsor, 1992, Brady & Brown)

102

Figure : alternative Methods of grouting cable bolts

The Rock bolting is applicable for smaller length of hole but where a longer length of hole is

drilled for bolting cable bolting is practiced. Longer length is followed to ascertain anchoring

of fractured rock at the back. This method has been developed during post pillar method of

stoping.

For cable bolting a 8 -10 m long hole is drilled at the back of the stope. A used wire rope with

open strands at the top is inserted into the hole along with a plastic breathing tube fastened to

103

the wire rope. Then a tube for pushing the mortar is placed at the lower portion of the hole

and the mouth is subsequently sealed. After this mortar is pumped into the hole and as the

hole is being filled the entrapped air is released through the breathing tube. Once the hole is

completely filled it is left for curing of the mortar. Cable bolts are used in addition to normal

rock bolts.

Back fill Materials:

The sources of fill are rock, gravel, river sand and mill failings. Previously hydraulic back

filling was carried out by means of river sand. In some of the Indian mines stopes were filled

by river sand against timber barricades lined with bamboo mats. However establishment of

milling facilities at the pit head and availability of mill tailings completely revolutionized the

filling system using back fill.

Many metallic ore deposits such as gold, copper, lead-zinc uranium etc. yield a large volume

of tailing from mill which can be suitably utilized for filling the stopes at the same time

solving the problem of their disposal. However depending on the nature of mineralization

much of the tailings from the mill sometimes upto 30% of the total is in the form of slime

which cannot be used as filling material.

Mill Tailings as fill Material :

A small fraction (2-5% by weight) of economic mineral is being recovered after crushing,

grinding and concentration of ore. The remaining rock mass having a wide range of size

distribution, is used for back-filling. Invariably the total tailing obtained from the

concentrating plant, undergoes certain processing before it is used as fill material. An

important aspect of the fill processing is maintenance of a suitable relation between the fill

quantity and quality, the one in general varying inversely with the other.

Relevant data are collected from a mine to determine the suitability of mill tailing for use as

backfill material. The underflow and overflow samples of mill tailings from hydro cyclones

are subjected to size analysis.

Size analysis fractions of a mill tailing.

The underflow has fairly a large portion of coarser fraction and the overflow contains mostly

slime and bulk of the water. The overflow contains the finer particles and rejected as slimes.

Example of fill material:

With a single stage hydro cyclone, about 50% by weight of ore milled is recovered as fill

material. Rubber lined hydro cyclones are used for recovery

(Micron size)

(%)(%) (%)

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The underflow of the hydro cyclones has a pulp density of 70% by weight, after proper

mixing with water, it is taken down the mine through bore holes by gravity action with solid

concentration of 60% by weight.

The most important characteristics of tailing sand suitable for hydraulic filling is its particle

size distribution, settling rate and compressibility (supporting ability). Both permeability and

settling rate are retarded by the presence of high proportion of slime. Discussed below are the

three important characteristics.

Fig. Modes of support of mine back fill

a. Support against the movement of the loose material from the wall rock

b. Local support ; c. Global support

Percolation rate: This depends mainly on shape, size and size distribution of particles. The

recovery of tailing sand varies from ore to ore. Finer the grinding in the mill more is the fines

which is to be rejected as slimes. In case of lower percolation rate it would be advisable to

allow longer time for drainage of water.

Settling rate :Most classified mill failings having an average size of about 50% settle pretty

fast. However with increasing percentage of slime ( which is sometimes due to increase the

sand recovery) settling takes longer time.

Supporting ability :Settled failing fill undergoes least amount of compression.

(a)

105

Fig. Simplified view of structure of a composite rock fill (Brady & Brown, Gonano 1977)

Cemented fill:

Another important development in filling has been the use of cemented tailing-sand fill.

Though initially developed for providing solid mucking floor in the stope its use has been

extended for filling the entire stope.

Scraping on ordinary hydraulically placed fill generally involves covering of the fill with a

tight-fitting timber floor of 50 mm thick planks in order to minimize dilution of ore with fill

and loss of fines. But it required additional cost for the timber and its placing and recovery

before and after filling. In certain mines, the timber floor is covered with a thin layer (50mm)

of fill in order to protect the timber from damage by blasted muck. In such cases, dilution of

ore with fill is unavoidable. In stopes where barren portions of muck are preferably left in the

stope before filling, use of a timber floor entails loss of timber. All these considerations have

led today to the adoption of cemented fill scraping floor in place of timber floors.

Scraping floors:

Compacted core

Porous, poorly

Cemented rockfill

Well

Cemented rockfill

Bedded,

Cemented sandfill

106

Scraping floors today use a cement: sand ratio varying from 1:6 to 1:15. while the floors with

a high cement content are only 100-150 mm thick, weaker cement floors have to be thicker.

Cemented tailing-sand floors are also ideally suited for stope mucking by rubber-tyred

loading machines. The strength of the floor has however, to be sufficient for these loaders.

Consolidated backfill:

The development of a consolidated backfill was primarily necessitated to provide a solid wall

when mining pillars in between previously mined and filled stopes. The normal practice to

prevent sand from adjacent filled stopes rilling down into the stope is either to leave a

retaining rib of ore against the adjacent stope which is not practicable except in low grade

ores or to build a suitable retaining timber barricade.

Properties of cemented fill:

Much laboratory work has been done on the characteristics of cemented fill and the factors

affecting them. The two main properties of a cemented fill are the permeability and strength.

The rate of development of strength is also important, particularly where a fast stoping cycle

is aimed at.

Permeability of cemented fill is basically determined by the particle size distribution of the

tailing-sand used in weak cement: sand mixtures, but rich (1:5) cement: sand mixtures

practically become impervious after curing.

Strength (normally uniaxial compressive strength is taken for purposes of comparison and

design) of cemented backfill depends on several factors such as (a) the cement percentage in

the backfill, (b) fineness of cement, (c) particle size distribution of the sand; (d) pulp density

when placing, (e) curing time, (f) temperature of setting etc.

Fig. 1 below give the variation of compressive strength with various cement to tailing-sand

ratios for various setting periods.

As is evident from the curves, the strength increases with cement content and time of setting,

though the rate of increase in strength falls sharply after a setting time of 28 days, normally

recommended for cement mortars. With the stope divided into sections so that ore breaking

goes on in one section at a time, it is possible to allow this setting time for the cemented

backfill, before another pour is made

The particle size distribution of the tailing sand affects the strength of the cemented backfill

to a fair extent. Laboratory studies reveal that classified tailings (with slimes removed) impart

a greater strength to the cemented backfill than unclassified tailings and that removal of the

finer fraction improved strength to a greater extent than the removal of the coarser fraction.

Pulp density is perhaps the most important factor in determining the strength of cemented

backfill. Fig. 2 illustrates the effect of pulp density on the strength of cemented backfill.

Apart from the strength, a higher pulp density helps in lesser segregation of the slurry and

causes less water to be handled in the system. However, a very high pulp density

substantially increases the apparent viscosity of the slurry and raises the critical velocity.

107

Fig. Strength variation with cement content

For successful implementation of any system, suitable design is necessary. Filling system

design for classified mill tailings some of the major factors are the hydraulic process,

mechanical aspects of classification system, the economic aspects and the operational

characteristics. Proper system design will render efficient rate of stowing without hindrance

which ultimately lead to higher productivity `of the mine in general and stopes in particular.

5. Vertical Crater Retreat method

VCR stoping is applicable in many cases where conventional shrink stoping is

feasible, although narrow orebody widths (less than about 3 m) may not be tractable. The

method is also particularly suitable for mining configurations in which sublevel development

is difficult or impossible. These geometric conditions arise frequently in pillar recovery

operations in massive ore bodies.

108

Figure Vertical Crater Retreat method

109

Figure Vertical Crater Retreat method

Figure Vertical Crater Retreat method

VCR Stoping Method:

One of the most recent methods to be adopted in underground metal mining is the vertical

crater retreat (VCR) mining which is now being employed in over large and medium scale

underground mines in different countries. The application of this new and revolution and

mining method has been possible only after down-the –hole drills were introduced to

underground mining operations. The method employs large diameter long holes of 152, 165

or 200 mm diameter and is based on the spherical charge technology (also known as crater-

blast technology) which is used to produce a series of craters in a horizontal plane, as a result

of blasting.

Mechanics of Crater Formation:

The method of VCR mining utilizes concentrated or spherical charges as opposed to

conventional cylindrical charges. A charge is considered to be spherical if its length-to-

diameter ratio does not exceed 6 to 1. Thus for a hole of 165 mm diameter, a slurry package

110

of 165 mm diameter and 990 mm length would form a spherical charge. The geometrical

configuration of a spherical charge limits its weight to approximately 35 kg in a 165 mm

hole. These spherical charges are placed in vertical or near-vertical parallel blast-holes at an

optimum distance from the bottom of the hole. The optimum distance (also called the depth

of burial) is defined as the distance from the free surface to the centre of gravity of the charge

and is so chosen that the maximum volume of rock is broken to an excellent fragmentation

size. When the charge is detonated, it produces a crater (surface cavity) in the surrounding

rock. As gravity works with the explosives breakage process and as the explosive energy in

spherical charges is used at optimum confinement conditions, the resultant crater depths

normally exceed the top of the explosives charge location and the muck produced is very –

well fragmented for an efficient handling.

The Stoping Method:

The VCR method requires large diameter holes, usually of 165 mm diameter, to be drilled in

a parallel pattern from a top drilling drive in the roe (called an over cut) down to an undercut

on the level below. When the drill pattern has been completed over the whole stoping block,

the bottom of each hole is blocked off and charged with ‘spherical’ slurry bags placed in the

hole at an optimum depth of burial. Horizontal slices of ore up to about 5 m thick, are then

blasted into the undercut. The ‘swell’ of broken ore is then drawn off (as in shrinkage

stoping) from draw points by LHD equipment, prior to the next blast being taken. After each

blast has been drawn off, the space between the top of the broken ore and the face of the

stope is measured which forms the basis for determining the thickness of the next slice to be

blasted.

Repeating this loading and blasting procedure, mining of the stone or pillar retreats in the

form of horizontal slices in a vertical upwards direction until the entire block is crater-

blasted.

The VCR method necessitates the use of water jel or aluminized slurry explosive having high

densities, high detonation velocities and high bulk strengths. ANFO, because of its low

density has not been used in VCR blasting, despite its attractive cost and safety

characteristics.

Several patterns of millisecond delays for blasting the slices of the ore body are used but the

preferred method isto first blast a burn cut out of the centre of the pattern while the remaining

holes are then blasted concentrically around the burn. This method gives each hole two free

faces into which it can break, laterally into the burn and downwards into the horizontal stope

back.

111

Figure: Schematic layout of VCR stoping

(a). mining primary stopes, and (b). mining secondary stopes (Harmin 2001)

Diameter, Length and Inclination of Blastholes

Most VCR mining has been done with hole diameters in the 152 or 165 mm diameter range.

Blastholes in the 200 mm range have been successfully used and can give good production

112

rates, although vibration considerations and the inability to execute larger drill patterns in

narrow widths can be the limiting factors.

The blast holes are usually about 50 m long and the efficiency of the stoping method is

largely dependent upon the drilling accuracy, since a poor configuration of holes produces a

substandard blast. In this regard, the method works best where the orebody dips at angles

greater than 700 since the static pressure of the drill string in an inclined hole is greatly

reduced.

Although some blastholes in excess of 90m length have been used, experience suggests that a

reduction in depth of 75m or even 60m can result in lower overall mining costs because the

higher development costs are then offset by improved results arising from greater drilling

accuracy.

Advantages:

VCR method has gained popularity both as a stoping method and for pillar extraction, in

conditions where suitable ore blocks are available and the rock mechanics aspects are

favourable. The VCR stopes have been used both as sublevel and shrinkage stopes. The

method has also been used in drop raising. The main advantages of this method include:

(i) Higher tonnage per day and lower stoping cost.

(ii) Lower development cost since it eliminates raise boring and slot-cutting.

(iii) Increased safety of operations because drilling and blasting are

carried out from above and there is no need for the miner to enter the actual stope.

(iv) Improvement in fragmentation (the method yields lowest powder

factor).

(v) Reduced labour requirements and drilling and charging time.

(vi) Reduced dilution and over break.

(vii) Elimination of up-hole drilling and up-hole loading of explosives.

Spherical Charge basis for VCR Method:

The term cratering is applied to the formation of a surface cavity in a material through the

action of detonating an explosive charge within the material. A crater blast is a blast when a

spherical charge is detonated beneath a surface that extends laterally in all directions beyond

the point where the surrounding material would be affected by the blast.

In analysing crater blasts, it has been found that there is a definite relation between the energy

of explosive and the volume of material that is affected by the blast and this relationship is

significantly affected by the placement of the charge.

Livingston has shown the importance of the shape of the charge in the breakage process. The

effect of the charge shape was demonstrated by detonating two equal charges of the same

explosive but of different shape in the same type of rock.

Table 1 Comparison of spherical and cylindrical charges

113

Spherical charge Cylindrical charge

Charge weight 4.5 kg 4.5 kg

Hole diameter 114 mm 67 mm

Diameter-to-length ratio 1:2.7 1:15

Volume of crater 4.4m3 1.1m3

Crater radius 1.7m 1.5m

The investigations on the effectiveness of the geometry of the explosive column have

reported that the spherical charge breaks a much greater volume of material than the

cylindrical charge.

(a)

(b)

Figure. Schematic of Spherical charge (a). side view (b). Plan view

114

The explosive’s contribution in blasting is to provide pressure. The forces generated by

pressure acting over a borehole’s surface area accomplish the necessary work to cause stress

conditions within the surrounding mass for fracture and displacement.

The explosion produces two distinct and separate pressures. The first is the detonation

pressure developed as the detonation front passes through the explosive charge. The

explosive’s detonation velocity directly affects the magnitude of this pressure. The value of

the detonation pressure is approximately proportional to the explosive’s density and its

detonation velocity squared. This pressure is applied at only a very short period of time

against the surrounding mass at a given section of charge length.

Livingston has found the relation,

Db = ∆ EW⅓

Where

db = the distance from surface to the centre of gravity of the charge i.e. depth

of burial

∆ = db/N dimension less number expressing the ratio of any depth of burial

compared with critical distance.

When db is such that the maximum volume if rock is broken to an excellent fragment size,

this burial is called optimum distance (do).

N - Critical distance at which breakage of surface above the spherical charge does not

exceed a specified limit.

N = EW⅓

Where

E – Strain energy feet

W – not of explosive charge

Borehole pressure dominates:

The second pressure that quickly follows the first is the borehole pressure produced by the

high temperature gases formed by the chemical reaction. The entire surface area of the

115

borehole where the explosive is contained will be exposed to a sustained loading condition. It

would be, therefore, the borehole pressure which dominates in the process of breaking the

rock.

Dynamic loading by borehole pressure in a cylindrical hole is predominantly directed

laterally, or radially outward from the borehole axis, with little or no force being directed

towards the charge ends.

The breakage mechanism of a spherical charge is quite different. The forces produced by a

spherical charge are directed radially outward from the centre in a uniform spherically

diverging action in all planes passing through the centre. It follows that the entire surface area

of the cavity confining the spherical charge receives all the detonation pressure, and the

borehole pressure.

It has been found that as long as the deviation from the true spherical charge (diameter =

length) is not greater than 1:6 diameter to length of charge ratio, the breakage mechanism and

the results are practically the same as that of a true spherical charge.

116

Spherical charged hole

Blast Hole Open Stoping:

Introduction of down-the-hole drilling equipment under-ground has revolutionized open

stoping and under suitable ground conditions blast hole open stoping has become a favoured

mining method all the world over.

117

(a). Plan view of the LDBH stope

(b). cross sectional view of the LDBH stope

Fig. First step in the opening of the slot.

118

(a). Plan view of the LDBH stope

(b). cross sectional view of the LDBH stope

Fig. Second step in the opening of the slot.

119

(a). Plan view of the LDBH stope

(b). cross sectional view of the LDBH stope

Fig. Third step in the opening of the slot.

120

Figure illustrates the blasthole stoping method adopted at Mufulira copper mine, Zambia.

Stope development consist of driving a 3.6x4 m gathering drives 15 m above the haulage

level (60 m below the upper haulage level) in the footwall of the ore body. Gathering cross-

cuts of the same cross-section are driven into the ore body from this gathering drive. A trough

drive connects the ends of the gathering cross-cuts. In ore bodies greater than 12 m wide two

parallel through drives are used one on the hanging-wall and the other on the footwall side in

order to reduce the size of the crown pillar and consequently the length of the holes required

to drill it, as well as to provide a good under cut. The gathering cross-cuts are connected to

the haulage level by box (chute) raises which can hold a train load (50t) of broken ore.

Fragmentation is good enough for transferring the broken ore through grizzlies installed on

top of the box raises.

The stopes are usually 40 m long with 12 m rib pillar though the stope length may vary

depending on the hangingwall conditions and the resulting dilution. The stope is undercut by

fans of holes of normal diameter (57 mm) drilled from the trough drives. Slotting is done at

the centre of the stope instead of against a rib pillar in order to prevent the weakening of the

latter by heavy blasting. The slot is started by boring a 1.8 m diameter slot raise in the

footwall. A mini slot raise is also driven up to the undercut level on the hanging-wall in order

to make the blasting of the slot fans from the trough drives easy. Parallel slot holes of 57 mm

diameter are then drilled from the drilling cross-cut above breaking the ore into the slot raise.

Drilling cross-cuts of 3.6 x 4 m size are driven from haulage level above towards the ore

body. Between the walls cuts are 9m apart centre to centre. 150 – 165 mm diameter hole are

drilled in two rows along the cross-cuts 0.75 m away from the sidewalls of the cross-cut. This

drilling pattern provides a burden of 3.5 m between rows within a cross-cut and of 5.5 m

between rows of adjacent cross-cuts. The spacing of holes in the two rows within a cross-cut

is varied between 5 and 6 m, one of the rows having larger spacing than the other.

Blasting of one complete cross-cut is done at a time. In order to minimize blasting vibration

each hole is charged in three decks with a 2 m sand stemming in between decks. Each deck is

initiated by a separate millisecond delay detonator so that no more than 200 kg of charge is

blasted on one delay. Each hole is first blocked at the bottom by a wooden block and then

provided with a 3 m toe stemming before placing of charge. There is a 3 m collar stemming.

The explosive used in ANFEX with pentolite booster.

Rib and chain pillars are drilled and blasted after the stope has been drawn out. While rib

pillar is drilled from the drilling cross-cut within the pillar in the same way as the stope, the

chain pillar is drilled from the haulage drive by 150 mm diameter upper holes not exceeding

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550 in inclination ( the limit for pneumatic charging of ANFEX into the hole). The fans are

spaced 6 m apart with a maximum toe burden of 5 m within a fan.

The blast hole method of stoping has resulted in a one-third cost saving as compared to

mechanized open stoping. The rate of drilling at present is 15 m/shift with an average

penetration rate of 3 m/h. Drilling tonnage is 60 t/m in stoping and 30 t /bag of ANFEX in

stoping and 55 t/bag in chain pillar blasting. Blasting fragmentation is excellent so that more

stopes are now using grizzley transfer than draw point loading as practiced earlier.

Drilling long large-diameter holes underground accurately is the key to the success of this

method of mining and underground down-the-hole drilling is essential. A HR 22 drill was

found inadequate. Robins 11 MD mobile rotary drill was found 3 times costlier than Ingersoll

Rand CMM or Atlas Copco ROC 306 DTH drills. The rotary drill could be competitive only

if it could be used at the higher penetration rate achievable by it, but at this penetration rate

the hole deviation was large. For tolerable hole deviation, the penetration rate had to be

brought down to the level achievable by the DTH drills. Hence the DTH was the obvious

choice. Of the two DTH drills mentioned ROC 306 was selected because of its ability to drill

upper holes required for drilling the chain pillars.

Blast hole stoping has been successfully used at Kolihan in India where a drill factor of 40

t/m is obtained as against 5 t/m in the conventional method. The tonnage per metre of

development increased from 140 t/m to 250 t/m. That was a 20-30% reduction in

development, 50% reduction in drilling time, 40% reduction in preproduction time and a 20-

30% reduction in overall stoping cost. Fragmentation was good with 20 t/kg of secondary

blasting. Ventilation is better, supervision concentrated and blasting more flexible.

6. Sublevel Caving

Ground Conditions

Ore strength: moderate to fairly strong, should be competent to stand without

support

Host rock strength: weak to strong, should becavable.

Deposit shape: tabular or massive

Deposit dip: steep(>60 degrees), can be flat if

the deposit is fairly thick.

Deposit size: large, extensive vertically

Ore grade: moderate

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Figure Sub-level caving details

Features

High production rate, large scale method

High recovery, high dilution

Suitable for full mechanization

Caving and subsidence occurs

Draw control important

High development costs

Some selectivity and flexibility

7. Block caving method

Ground Conditions

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Ore strength: weak to strong, must be fracturedor jointed and cave freely

Host rock strength: weak-moderate, similar toore in characteristics

Deposit shape: massive or thick tabular, fairlyregular

Deposit dip: steep(>60 degrees or vertical)

Deposit size: very large

Ore grade: low, uniform

Figure Block Caving

Features

High productivity, low mining cost (comparable to open pitmining)

Large scale method, high production rates

High recovery and potentially high dilution

Rock breakage by caving – no blasting costs

Large scale caving and subsidence, wholesale damage to surface

Good draw control essential

Slow, extensive and costly development

Highly mechanised

Inflexible

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Problems

Q. Define dilution of ore in stoping operations and discusses the effects of dilution?

Answer:

Dilution is defined as the low grade (waste or backfill) material which comes into an ore

stream, reducing its value. Ore loss refers to any unrecoverable economic ore left inside a

stope (broken, in place as pillars or not properly blasted at the boundaries), or to any valuable

ore not recovered by the mineral processing system. The detrimental impact of dilution to the

economics of the mining industry is well realised. Waste rock dilution and ore loss exist

during geological modelling and evaluation, decisions regarding cut-off grade, design of the

mining method, stoping and ore concentrating.

Dilution is a source of direct cost as waste rock or backfill material is blasted, mucked,

transported, crushed, hoisted, processed and stored as tailings. Dilution is also a source of

indirect cost as the dilution material may adversely affect the metal recoveries and

concentrate grades. A lost opportunity may result from directing resources at handling waste

(as opposed to ore) for the mill feed. Furthermore, ore processing facilities will be engaged

for material which contributes very little to final useful metal production. In most cases,

mining and milling capacity is limited; this capacity is affected by the displacement of ore by

waste within the overall mining and processing facilities.

(Source : Ernesto Villaescusa)

Q Define dilution and describe the classification of dilution?

Answer

Dilution is always defined and quantified with respect to an idealized (planned) stope

boundary. In order to quantify dilution, an orebody must be properly delineated and the

extracted volumes must be effectively measured.

Dilution can be divided into three general categories, namely; internal, external and ore loses

(See Figure below). Internal dilution usually refers to the low-grade material contained within

the boundaries of an extracted stope. It can be caused by insufficient internal delineation of

waste pockets within an orebody. It is also occur in situations where the mining method dictates

a minimum width of extraction. External dilution refers to the waste material that comes into

the ore stream from sources located outside the planned stope boundaries. Low grade material

from stope wall overbreak, contamination from backfill, and mucking of waste from stope

floors are typical examples of external dilution. Ore loss refers to the economical material that

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is left in place within the boundaries of a planned stope. Planned ore diaphragms (ore skins),

unbroken stope areas due to insufficient blast breakage, non-recoverable pillars left to arrest

stope wall instability and insufficient mucking of broken ore within stope floors are typical

examples of ore loss.

Geological dilution refers to the waste rock or ore-losses incurred during the exploration and

orebody delineation stages, where only an estimated model of the orebody can be made. A

geological model is based on limited information, and is unlikely to coincide exactly with the

real orebody; therefore the delineated orebody boundaries are likely to exclude ore and also

to include waste.

Figure classification of dilution (Source : Ernesto Villaescusa)

Question:

Briefly describe the special stoping methods ?

Blast hole technique of sublevel stoping

This method is a modification of sub-level stoping where ring drilling is replaced by

longdiameter parallel holes and also multiple sub levels are eliminated. This utilizes 170mm

large diameter DTH drills and is more efficient than the ring or fan drilling while the

development, drilling factor and powder factor are concerned. It results in consistently

acceptable fragmentation. A hole is loaded with alternating decks of explosive and inert

material. The decision to load the hole with a column charge or alternatively with a decked

charge is based on the consideration of allowable powder factor, hardness of ore,

fragmentation requirement and blast vibration.

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Overhand and underhand stoping operations

The overhand and underhand stoping operations are based on direction of blasting. Any

stoping method can therefore be classified as an underhand method or an overhand method

depending upon the direction of advance of stoping These methods are confined to narrow

steeply dipping orebodies with competent H/W and F/W and requires free flow of blasted ore

in the stope. These are highly labour intensive methods. The main difference between the two

is the commencement of stoping operation from lower to upper level for over hand and upper

to lower level for underhand. For low dipping orebodies under suitable circumstances can be

mined by breast stoping and inclined room and pillar method.

Open overhand and underhand stoping:

Open stoping methods are designed for widely different conditions of dip, width of deposit,

character of ground and grade of ore. The following principles may apply.

Low cost mining by open stoping is sometimes possible through sacrifice of part of

deposit.

Underground stoping of ore is possible in flat deposits but only to a limited extent.

Use of open stoping usually presupposes strong ore and strong walls.

Methods are usually limited to tabular deposits with regular well defined wall.

The underhand and overhand stoping are confined to narrow steeply dipping ore bodies

where both hanging wall, footwall and the ore are strong and require little or no support. Free

flow of the blasted ore should be there. As mechanization is not possible they are low

tonnage and labour intensive method.

Stope development is commenced by driving a series of main levels normally in orebody.

Maximizing the level interval reduces the development to stoping ratio. Benefits of minimal

level spacing is the possibility of close sampling of the vein. In cases where closure will

occur the use of crown and sill pillars is practiced. Levels are connected by raises at suitable

intervals. Further development is in the form of stope drive if a pillar is to be left between

main level, and the stope

Underhand stoping method:

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Stoping commence from upper stope drive outwards from the raise connection. Stoping

action is done in a series of benches approximately 2m high and 1.5 m wide. An overall

stoping line at approximately 600 is maintained. The benches are normally drilled vertically

downward and blasted in sequence starting at the bottom of the face. As benching continues

the lower level is reached and stope face retreats thus permitting more than one boxhole to be

used for on recovery.

Fig Underhand stoping operation

Overhand stoping method:

It is similar in basic concept to the Underhand stoping. However, in this case the stoping

operation commences from the lower footwall drive. Again the stope may be commenced

above a pillar through which box hole raises are driven. The miner is supported during the

stoping operation on platforms made on stull timbers which are extended upwards as stoping

progresses. Drilling is done by stopers.

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Fig. Overhand stoping operations

Breast stoping:

The term breast stoping denotes primarily a method of breaking ground by advancing early

horizontally a vertical face or breast of ore. This method is used in low dipping ore bodies

and mined as open stoping. Typical breast stoping ore bodies are normally less than 2.5 meter

in thickness although this figure could be greater or less depending on ground conditions and

the support system chosen. This is a low cost method which usually sacrifices some ore in

permanent pillars hence especially suited to low grade ore where high extraction is not of

primary importance.

Ore is broken by slightly inclined holes drilled in a vertical face of considerable lateral area

which is being advanced in a horizontal direction. Handling of ore is by shoveling or using

scrapers and drawn to lower level.

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Fig. Breast stope

(a). Schematic diagram of inclined room and pillar

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Fig. Inclined Room-and-Pillar stope- Ore handling by a Scraper system

Incline Room and Pillar :

This method is similar to breast stoping in so far as drilling, blasting and mucking were

concerned. Stope timbering has been eliminated completely and instead 1.5 m long rock bolts

at 1.2m x 1.2m spacing in a systematic pattern are used as roof support. Chute and crown

pillars of 5m down the dip and rib pillars of 3m (for 10w span) and 4m ( for 15m span) are

left between the rooms.

After mining the first two meters of the lode width along the HW contact the ore left in the

footwall are stripped out in stages by drilling 1.5 to 2m long holes. Since the HW is rock

bolted the work in the stope can be done freely. Lode width upto 6m and inclination more

than 500can be mined by this method.

The field trips to the mines and other geological sites are essential part of the learning of this

module. The regular geological field trips involve identification of different rock types as

well as other exercises related to plotting sterographic projection for various rock

discontinuities. These exercises help them recognize joints, number of joint sets, condition

and extent of joints and it will help the students understanding of the importance of structural

discontinuities and know their impact on the rock mass characterization.

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Q. Explain the economic considerations in stoping operation.

Economic considerations: For an investment in a new mining venture there are a

number of potential pitfalls that the investor may be exposed to.The economic feasibility of a

mining project is controlled by the reserve, the grade, the target production and the expected

cycle life of a mine and the present value of the mineral. In order to conduct a suitable and

sufficient economic evaluation the investor needs to establish the likely price of the mineral

product as well as future trends envisaged in the price of the minerals of economic interest in

the ore body. This is an extremely important aspect of the evaluation as mines are normally

price takers and not price makers.

The actual configuration of the ore body will also play a significant role in the evaluation

process. Some ore bodies occur as a massive homogeneous deposit that makes mining easier

while other may occur in numerous veins or lenses that could be flat or vertical and may be

geographically distributed in the lease area. These have an impact on the cost to create a

suitable access to the ore body and the building of the extraction plant.

At every stage there is an uncertainty and hence the evaluation and selection of mining

methods is a process that requires both the knowledge of mining methods as well as a strong

working knowledge of the methods of cost estimation. Knowledge of both areas is necessary

because the work of providing cost comparisons require that the cost estimator be familiar

with mining methods to provide accurate cost predictions. The choice of a mining method

and the decision whether to pursue the development of a mining property are closely

interrelated.

The purpose of the mine design, as it relates to estimating costs, is to determine equipment,

labour, and supply requirements both for preproduction development and daily operations.

The extent to which the mine is designed is important.

The economic feasibility of an ore deposit is dependent upon the following basic parameters:

(a). Minable tons(Reserve)

(b). Ore body grade

(c). Mineral value

(d). Production rate (output per unit time)

(e). Mine life

(f). capital cost

(g). Operating costs

Minable tons/ Mineable Reserve is those parts of the ore body, both economic and

uneconomic, that are extracted during the normal course of mining.

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Mineral Reserve is the economically mineable part of a Measured or Indicated Mineral

Resource demonstrated by at least a preliminary feasibility study. This study must include

adequate information on mining, processing, metallurgical, economic, and other relevant

factors that demonstrate (at the time of reporting) that economic extraction can be justified. A

mineral reserve includes diluting materials and allowances for losses that may occur when the

material is mined.

Figure Valuation methods depending on the stage of development on the mineral property

Orebody grade, Mine dilution and recovery factors

An ore is a type of rock that contains minerals with important elements including metals.The

grade or concentration of an ore mineral, or metal, as well as its form of occurrence, will

directly affect the costs associated with mining the ore. The cost of extraction must thus be

weighed against the metal value contained in the rock to determine what ore can be processed

and what ore is of too low a grade to be worth mining. Metal ores are generally oxides,

sulfides, silicates, or "native" metals (such as native copper) that are not commonly

concentrated in the Earth's crust or "noble" metals (not usually forming compounds) such as

gold. The ores must be processed to extract the metals of interest from the waste rock and

from the ore minerals. Ore bodies are formed by a variety of geological processes. The

process of ore formation is called ore genesis.

Mining seldom recovers all resource present in an ore deposit. The amount of ore actually

extracted from a deposit is referred to as the recovery factor and is expressed as a percent. In

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addition, a certain amount of waste is usually mixed in with the ore during mining. This

waste mixed in as ore is called dilution and is usually expressed as a dilution factor (in %).

Both recovery and dilution vary with each ore body, but tend to be within a similar range for

each mining method. Table below summarizes the assumed dilution and recovery factors

used for the mine models and reflects values commonly encountered when these mining

methods are applied.

Table Dilution and recovery factors

Mining method Dilution

Factor %

Recovery

factor %

Block caving 15 95

Cut-and-fill 5 85

Room-and-pillar 5 185

Shrinkage 10 90

Sublevel longhole 15 85

Vertical crater

retreat

10

Mineral value

The price of the mineral at a given time is one of the important factors in the selection of a

suitable mining method because it decides the viability of the proposed project. In order to

conduct a suitable and sufficient economic evaluation the investor needs to establish the

likely price of the mineral product as well as future trends envisaged in the price of the

minerals of economic interest in the ore body.

This is an extremely important aspect of the evaluation as mines are normally price takers

and not price makers.

Production Rate and Mine Life: Given a known ore reserve tonnage, the life and daily

capacity for a typical mining operation can be determined. Taylor developed an equation

commonly used in prefeasibility studies to determine mine life, known as Taylor's rule. Based

on this rule, the basic equation for C (capacity of ore production in st/d) is:

𝐶 =𝑇

𝐿 × 𝑑𝑝𝑦

Where:

L = mine life in years

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T = Total tonnage factor of ore to be mined

dpy = Operating days/year

𝑇 = 𝑟𝑡 × 𝑟𝑓 × (1 + 𝑑𝑓)

Where:

rt = total deposit (reserve) in tonnage (st)

rf = recovery factor for the particular mining method

df = dilution factor

Substituting for L using Taylors rule

𝐿 = 0.2 × 𝑇0.25

We can determine the daily mining capacity(output) by the following expressions:

𝐶1 =𝑇

350 × 𝐿=

𝑇0.75

70

Or

𝐶2 =𝑇

260 × 𝐿=

𝑇0.75

52

Where C1 = mine capacity in st/d for 350 days/y and 7 days/week

C2= mine capacity in st/d for 260 days/y and 5 days/week.

The Life Cycle of a mine

Figure below demonstrates the life cycle of a mining share, which shows how the share price

behaves depending on the stage of the mining project. At more mature stages of the project

the risk goes down and the share price goes up.

Mining is a depleting business – “the more you mine, the less you have left to mine and

without exploration, mining will cease very rapidly. The mining companies know they need

access to good exploration projects and, more importantly, good exploration teams.”

Therefore it is important that a company’s management has the ability to generate new

exploration projects. The figures below illustrates the life cycle of a mine investment and

share value.

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Figure The life cycle of a mining share (US Global Research)

Capital cost: Capital costs are based on actual equipment list prices in most cases. An

additional cost is applied to all equipment purchase costs for freight. Underground capital

costs were determined by the amount of development necessary for an underground mine of

the size and type under consideration to begin operating at design capacity.

Operating cost: Operating costs are based on daily capacity (short-ton/d) and are expressed

in rupees per short ton (Rs/st). All the underground models are based on st/d of production,

and costs are in Rs./st.

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Figure Underground mining capital and operating costs (Thomas W. Camm,1989)

137

Figure Underground mining capital and operating costs (Thomas W. Camm, 1989)

Method selection plays an integral role in these considerations since it impacts all factors

except mineral value. As a result, proper extraction method design dictates a project’s profit

margin, and in this sense, mineral value influences the mining method.

An ore body’s mineable inventory is a reflection of the tons and grade that can be mines at a

desirable profit. The mining method will significantly influence this inventory by affecting

selectivity.

For example: Open cut-and-fill stoping offers a high degree of extraction control and will

optimize the mineral content of every ton mined. Unfortunately, selective methods generate

higher operating costs because they are more labour intensive and consequently less

productive than bulk methods. This increase cost will often diminish the benefits of

optimizing mined ore grade.

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(1 ton=0.9072t, 1

oz/ton=31.25g/t)

Figure Comparative unit cost in US$ of productivity in VCR, Mechanised cut-and-

Fill and Open cut-and-fill methods of mining

The above figure illustrates the comparative unit cost in US$ of productivity in VCR,

Mechanised cut-and-Fill and Open cut-and-fill methods of mining. While open cut-and-fill

allowed 4% increase in extracted grade over VCR, the 55% lower productivity translated to

36%higher operating costs and resulted in an uneconomic method. Therefore the mechanized

cut-and-fill was specially introduced to replace open cut-and-fill as a relatively selective

method that employs higher mechanization to reduce operating costs and boost productivity.

Table Comparison of relative direct costs of various mining methods

Mining method Relative cost

Block caving 1.0

Room-and-Pillar 1.2

Sublevel stoping 1.3

Sublevel caving 1.5

VCR 4.3

Mechanized cut-and-Fill 4.5

Shrinkage stoping 6.7

Conventional Cut-and-Fill 9.7

Normally, a method is chosen that generates a minable inventory to sustain consistent

profitable cash flow for the longest period of time. This allows for full project capital

recovery and provides cash flow to use for exploration and development of additional

reserves. A factor such as market price, however can adversely affect this philosophy.

Commodities such as strategic metals that are subject to drastic price fluctuations, often

dictate initial high grading of an ore body to reduce capital recovery time. This means to

mitigate the risk of market volatility, and if the remaining mineral inventory can still be

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mined profitably, it may be the most profitable economic scenario. This is true particularly if

loans are involved and discounted cash-flow considerations include service of a debt. This

practice usually has the inherent disadvantages of sacrificing lower-grade reserves that could

otherwise have been blended into a consistent economic grade resulting in the ultimate

extraction of more mineral units. Essentially, the goal is to generate the optimal mix between

quickest return of investment and highest return on investment.

Choice of a mining method has a significant impact in capital requirements and revenue-

generation lead time. Some mining methods, such as block caving, are particularly

development intensive and, as a result, require more preproduction capital expenditure and a

longer lead time prior to revenue generation.

Cut-off grade is a dynamic number affected by commodity value and cost to produce the

product. In this respect, each mining method will generate a unique cut-off grade. Calculation

of this cut-off when choosing a mining method should be based on all operating costs

incurred to produce the commodity. This corresponds to the active cut-off grade utilized in a

producing mine.

Q. Why mining companies are different compared to any other traditional venture?

Answer

Mining is a depleting business – “the more you mine, the less you have left to mine and

without exploration, mining will cease very rapidly. The mining companies know they need

access to good exploration projects and, more importantly, good exploration teams.”

Therefore it is important that a company’s management has the ability to generate new

exploration projects. The figures below illustrates the life cycle of a mine investment and

share value.

Figure The life cycle of a mining share (US Global Research)

140

The prediction of the value of a mining company is a complex matter. Variousmethods are

available to estimate a company’s value but many are not useful orapplicable. The reason is

the specific nature of mining industry. Aside from the usual financing risk in the case of

mining producers, and financing and “finding” risk in the case of pure exploration

companies, there are price cyclicality, on going changes in operating and capital cost

structures, stock market vagaries, and volatility in circumstances. Consequently, even

traditional methods such as Discounted Cash Flow, Relative Multiples or Real Options

cannot be applied without some adjustments and demarcations. For example, cash flow or

earnings based valuation methodologies may not be relevant for the valuation of a mining

exploration company that has no production assets or revenues, neither operating cash flow or

earnings.

Q. What are the Characteristics of precious and industrial metals and give a brief

account of the classification of metals.

Answer:

Characteristics of precious and industrial metals

All mining activities take place within the Earth’s crust, about the top 7-35 km of the solid

matter comprising the bulk of the planet. The distribution of metals within the crust can be

seen by the differences in the types of rock which it contains: limestone, granite, sandstone or

basalt. Nevertheless, these different rock types are generally of uniform composition and

further concentrations need to occur in order to produce concentrations of material which can

be mined and sold at a profit. Therefore, the importance of the concentration factor

in determining the value of mining company should not be undervalued. A company with a

lower grade of ore will have to process more rock, possibly at greater cost in order to obtain a

given amount of economically valuable material.

Metals classification is presented in the Figure below. The precious metals are relatively rare

they are widely traded and are thought of as financial security in times of war or financial

crisis. The base metals have wide range of applications throughout industry and could be

thought of as the industrial metals.

The minor metals are produced very often both as by-products of the extraction of the major

metals or are required for specific applications and are therefore produced sometimes in small

quantities from primary deposits. It can happened that, if new producer brings a low cost

mine into production or if there is a massive increase in demand due to the discovery of a

new application, prices swing widely.

141

A lower grade gold ore would contain something like 5 grams per tonne (5 parts per million).

So, gold ore needs to be concentrated by about 1,000 times above its average dispersion to

become viable for gold mining.

Figure Classification of Metals

Q3.What are the special features of metals and mining companies?

Answer:

Mining and metals industries are highly cyclical in nature. The valuation of a mineral

industry is different from other traditional companies because of the swings in the demand for

the mineral over a period of time. There are two cycles in the process: one is commodity

price and/ or the other one is economic cycle. Commodity companies can determine the price

of commodity by changing amount of their production. Because of big changes in the prices

of mining company’s products, they are characterized by highly volatile earnings and cash

flows over a number of years.

The resulting valuation will greatly depend on where in the cycle (economic or commodity

price) the company stands. When commodity prices, say the metal prices, are in upswing or

in boom phase, all producers of this commodity benefit, whereas an extended economic

downturn or a lengthy phase of a low commodity prices burdens operators, even the best

companies in the business. Consequently, commodity companies are exposed to cyclical risk

over which they have little control.

The value of the commodity company is not only affected by the price of the commodity but

also by the expected volatility in that price. Commodity companies experience far greater

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price volatility than manufacturing companies or service companies do. This leads again to

volatile revenues, earnings and cash flows of the commodity company.

The other special feature is high fixed cost, thus commodity companies may have to

keep mines operating even during low points in price cycles. The reasons for this are

prohibitive costs of shutting down and reopening operations. Indeed, in a worst case scenario

such events could even force the mine to close and put the company into liquidation before

the exhaustion of its reserves.

It is important to mention that for metals and mining firms to get started, large

infrastructure investments are needed. It has led to the fact that many of these companies are

significant users of debt financing. Because of this, the volatility in operating in come that is

referred to earlier manifests itself in even greater swings in net income. Also when a

commodity company seek opportunities to extend its existence beyond the life of its reported

reserves in new areas, one of the main financing will be debt financing consequently, metals

and mining companies have high volatility in equity values and debt ratios.

Next, the mining industry has long lead times (e.g. ordering equipment like a mill)to

bring on new capacity. The mine development process is very specific and can typically take

5-10 years or more. Thus, most of these projects will begin their operations after many years.

The consequence of long lead times is a high risk for mining projects. Mining projects may

have many different risks, depending on the specific situation of the project. The most serious

risks include:

financing risk: equity (can funds be raised in the market), debt (interest rate, requirement of hedging by the lenders

permitting risk

Issues associated with geology (size and grade of the mineable portion of the orebody) and how the deposit can be economically mined.

Metallurgy (often underestimated – how much of the metal can be recovered, what is the preferred recovery method; are there any impurities or associated minerals that could affect this?)

Economics (metal markets and their forecast behavior, transportation costs, interest rates)

Country risk:

political risk (government stability, taxation instability, laws, environmental policy)

economic risk (currency stability, foreign exchange restrictions). Metals prices and metals’ stock performance are strongly correlated to exchange rates and particularly to the US dollar. This is primarily because over 70%of materials production comes from outside US dollar-denominated regions. As the dollar strengthens/weakens it alters the production

economics of suppliers and consumers.

143

Geographic risk (transportation, climate)

social risk (corruption, availability of workers and local labour laws, ethnic or religious differences within the indigenous population)

Lastly, earth has finite quantity of natural resources; therefore metals and mining is a finite

business. Mineral deposits contain a certain amount of ore and when that ore is mined out the

deposit is depleted, no matter what one does or wishes. The longevity of a commodity

company depends consequently onastute acquisitions, successful exploration, and/or a range

of non-mining or downstream businesses

When valuing commodity companies, scarcity of resources will play a role in what our

forecasts of future commodity prices will be and may also operate as a constraint of assuming

Q. Describe the mining method selection criteria

Answer: Underground mining method selection criteria

Evaluation

Parameters

Considerations

Geotechnical Lithological

Ground water

Geophysics

Ore genesis

Mineral

Occurrence

Continuity of ore zones within mineralised strata

Occurrence of mineral within ore zone(geological grade)

Economic mineral occurrence within ore zone (mining grade)

Ore body

Configuration

Dip

Plunge

Size

Shape

Safety/ regulatory Labour intensity of method

Degree of mechanization

Ventilation requirements

Ground support requirements

Dust controls

Noise controls

Gas controls

Environmental Subsidence potential

Ground water contamination

Noise controls

Air quality controls

Labour/Political Costs and influences.

1. Geotechnical evaluation:

Geotechnical considerations in method selection include an evaluation of lithology, groundwater,

and ore genesis of the deposit. This analysis should occur concurrent with the exploration drilling

144

phase and must include an evaluation of the ore zone hanging wall and footwall host formations and

general surface topography. As a result, a percentage of drilling must provide cores.

(a). Lithology: Important structural features like faults, folds, joints ets

could affect the integrity of both ore and host formations.

(b). Ground water: Important information like water levels within the

formation, permeability of the formation, initial flows, and sustained

flows.

(c). Geophysics: Basic data for evaluation includes tensile strength,

compressive strength, modulus of elasticity, Poisson’s ratio, angle of

internal friction and cohesion. Additionally the information regarding

the in-situ stress condition is also useful.

(d). Ore Genesis: This information is important in the sense that epigenetic

vein deposit typically contains high-grade ore shoots. Sedimentary

syngenetic deposits, unless they have undergone regional

metamorphism, are usually structurally incompetent.

2. Mineral Occurrence

The spatial distribution of mineral within the deposit can significantly impact the choice of

mining method. There is actually two fold considerations which constitutes continuity of ore

zones within the mineralised strata and the occurrence of minerals within the ore zone.

Ore bodies that occur as concentrated chutes or loads require aa very selective mining method.

This provides a significant grade control throughout the mining cycle.\ to ensure that dilution if

minimized.

The grade of a deposit is a reflection of mineral units entrained within the delineated tonnage.

This is normally termed as both geological grade and mining grade. The mining grade is always

higher because it is profit controlled. It will often influence the mining method because deposits

with a high-grade ore can often be mined at greatest profit by a selective method that ignores

lower grade reserves even though they are sometimes economical.

3. Ore body configuration

The physical parameters of an ore body will often preclude use of many mining methods.

Orientation considerations such as dip, plunge, and strike along with the size and shape of the

ore body are key evaluation factors. Many methods such as shrinkage, open stoping, sublevel

open stoping, VCR mining, sublevel caving, depend upon gravity ore flow to extraction points.

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Plunge is the vertical angle component along strike and it affects the number of mineral units

per vertical and horizontal linear distance. In this respect, plunge has a significant impact on the

amount of development required for each ton of ore.

The strike or azimuth of an ore body in the long dimension can affect a mining method by its

length. Open voids have a critical geophysical length-to-width parameter beyond which they are

no longer stable.

Open stoping , shrinkage stoping, or VCR mining that creates large openings are often designed

in a series of smaller panels to avoid hanging wall failure in deposits with a long strike

dimension. Figure below illustrates this point.

Figure Transverse panel mining along strike

4. Safety/ Regulatory factors

Health and safety of personnel is of paramount importance in the selection of any mining

method. Consequently a number of safety considerations need to be included in any mining

method evaluation.

An important selection of mining method can create serious ground-control related safety

hazards. For example, utilization of block caving in an orebody with poor natural caving

characteristics will create bridging problems. Any attempt to dislodge hang-ups expose

personnel and equipment to unstable ground, and eventual failure can result in an air blast.

Therefore all possible safety risks which could be created as a result of a method’s application

must be anticipated in advance.

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5. Environmental Factors

Most of the underground mining methods, except caving methods, have the distinct advantage

of creating a minimal disturbance to the environment. Any mining method that could result in

subsidence should be avoided.

Supported methods must ne evaluated depending upon ground competency.

Dust or gas emissions to surface air are limited to ventilation exhaust discharge location and

hence they should be properly located.

6. Labour and political issues

The choice of any mining method is influenced by the availability and cost of labour.

Political climate in many countries can result in relatively unstable governments. Significant

risks associated with these conditions favour a mining method that minimizes capital outlay

and provides the quickest return on investment. In these cases, mitigation of risk on

investment becomes a predominant factor in the choice of a mining method.

Unit Problems

Q Describe in details the role of geo-physics for mining method selection?

Answer: The details of geo-physics includes:

1. The geomechanical details such the physical and mechanical properties of the rock

cores are evaluated on the cores obtained from the exploratory drilling.

2. In situ stress measurements provide a quantification of existing stress within the

rock.

3. A series of measurements in multi orientations should be made on laboratory core

samples as well as in situ tests to (a). Account for planes of weakness such as

jointing or foliation and (b). Measure in situ stresses in all dimensions.

All the generated data is used to prepare computer models to project the areas of high stress

concentrations and possible rock displacement magnitudes during the mining sequence. This

will allow optimization of a mining method to the unique geophysical characteristics of the

deposit.

For example:

A method which creates large voids such as open toping, VCR mining, would probably be a

poor selection if high stress concentrations exist or would be generated in the hanging wall.

Subsequent failure would severely dilute ore and possibly threaten surrounding mine activity.

Both the open stoping and VCR stoping might be feasible methods if maximum vertical stope

height were reduced, stope length along the strike is shortened, or cable bolts added to the

hanging wall.

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Q Explain the planning for the spacing of excavations for various conditions.

Answer:

Spacing of excavations:

The following rules are based on the theory of stress concentrations around underground

openings and the interaction of those stress concentrations. The usefulness of these guidelines

has been borne out by experience obtained underground. Stress interaction between

excavations can obviously be controlled by an increase on the installed support, but costs will

also increase significantly. If there is adequate available space, it is generally more cost

effective to limit stress interaction between excavations.

Flat Development

(a). Square cross section(figure given below)

Spaced horizontally at three times the combined width of the excavations.

Spaced vertically at three times the width of the smaller excavation, provided

that the area of the larger excavation is less than four times the area of the

smaller opening

(b) Rectangular cross section (figure Below)

Spaced horizontally at three times the combined maximum

dimensions of the excavations.

Spaced vertically at three times the maximum dimension of the

smaller excavation provided that the height-to-width ratio of either

excavation does not exceed 2:1 or 1:2.

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(c) Circular cross section(figure below)

Spaced horizontally at three times the diameter of the larger

excavation.

Spaced vertically at three times the diameter of the smaller

excavation is less than four times the area of the smaller

excavation.

Vertical Development (eg shaft)

Square cross section at three times the combined widths of the

excavation.

Rectangular cross section at three times the combined diagonal

dimensions of the excavations

Circular cross section at three times the diameter of the larger

excavation.

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Q Explain with, suitable mining methods, the influence of economic considerations

on the selection of a mining method

Answer:

The economic feasibility of an ore deposit is dependent upon the following basic parameters:

(a). Minable tons

(b). Ore body grade

(c). Mineral value

(d). capital cost

(e). Operating costs

Method selection plays an integral role in these considerations since it impacts all factors

except mineral value. As a result, proper extraction method design dictates a project’s profit

margin, and in this sense, mineral value influences the mining method.

An ore body’s mineable inventory is a reflection of the tons and grade that can be mines at a

desirable profit. The mining method will significantly influence this inventory by affecting

selectivity.

For example: Open cut-and-fill stoping offers a high degree of extraction control and will

optimize the mineral content of every ton mined. Unfortunately, selective methods generate

higher operating costs because they are more labour intensive and consequently less

productive than bulk methods. This increase cost will often diminish the benefits of

optimizing mined ore grade.

(1 ton=0.9072t, 1 oz/ton=31.25g/t)

Figure Comparative unit cost in US$ of productivity in VCR, Mechanised cut-and-

Fill and Open cut-and-fill methods of mining

The above figure illustrates the comparative unit cost in US$ of productivity in VCR,

Mechanised cut-and-Fill and Open cut-and-fill methods of mining. While open cut-and-fill

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allowed 4% increase in extracted grade over VCR, the 55% lower productivity translated to

36%higher operating costs and resulted in an uneconomic method. Therefore the mechanized

cut-and-fill was specially introduced to replace open cut-and-fill as a relatively selective

method that employs higher mechanization to reduce operating costs and boost productivity.

Mining method Relative cost

Block caving 1.0

Room-and-Pillar 1.2

Sublevel stoping 1.3

Sublevel caving 1.5

VCR 4.3

Mechanized cut-and-Fill 4.5

Shrinkage stoping 6.7

Conventional Cut-and-Fill 9.7

Q Draw a comparative table between all the underground hard rock mining

methods in terms of their advantages and disadvantages.

Answer:

Mining

Method

Advantage Disadvantage

Sublevel

Caving

High degree of mechanization is

possible

Good selectivity can aid in grade

control it ore is at near vertical dip

High tonnage and productivity are

possible

High initial investment

Potential for dilution if ore body thin

or near horizontal

Black Caving Can be very cost effective

High production can be achieved

Grade control through draw-point

monitoring an asset

Surface subsidence

Big development effort

High capital cost

Room-and-Pill

ar

mining

High degree of mechanization

Possible with excellent

Productivities

Flexible and safe

Good grade control

Ground movement can become

onerous

Capital intensive

Ore left in pillars

Sub-level stopi

ng, blasthole st

oping, VCR st

oping

Easily mechanised

High Productivities

Large equipment cab be utilized

High capital investment

Grade control can be a problem

Strong engineering and technical

support required

Shrinkage stop

ing

Small stopes can be mined

Minimal developmental costs

Simple drilling and mucking

equipment

Broken ore required as fill material

Grade control can be difficult

Not a high tonnage or productive

Method

Cut-and-Fill Ground movement minimized

Dilution can be controlled easily

Compatible with non-filling

Labour intensive

Difficulty in ventilation

Can be costly and hence high grade

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Methods

Undercut-and-Fill techniques can

be utilized

Ore is required

Mill required for tailings fill material

Q What do you understand by the term ‘Mineral resource’ and ‘Mineral Reserve’.

Answer:

Resource and Reserve

For any mineral property, asset value is the extractable mineral resources located under the

earth’s surface and the invested capital is used mainly to extract this mineral resource. In

order to perform a fundamental valuation of a mining company the amount of mineral

reserves must be estimated. The definitions of Mineral Reserve and Mineral Resource are

given below:

Mineral Resource is a concentration or occurrence of material of intrinsic economic interest

in or on the Earth’s crust in such form and quantity that there are reasonable prospects for

eventual economic extraction. Portions of a deposit that do not have reasonable prospects for

eventual economic extraction should not be included in a Mineral Resource. The location,

quantity, grade, geological characteristics and continuity of a Mineral Resource are known,

estimated or interpreted from specific geological evidence and knowledge

Mineral Reserve is the economically mineable part of a Measured or Indicated Mineral

Resource demonstrated by at least a preliminary feasibility study. This study must include

adequate information on mining, processing, metallurgical, economic, and other relevant

factors that demonstrate (at the time of reporting) that economic extraction can be justified. A

mineral reserve includes diluting materials and allowances for losses that may occur when the

material is mined

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Figure Relationship between Mineral resource and reserve

Learning Strategy

This is completely a theoretical module and any exposure to the hard rock mines will enable a

student to comprehend the terms and definitions presented in the module.

References

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