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DR GIRI PRASAD REMINISETTY, profile picture
DR GIRI PRASAD REMINISETTY

1. India produces many important minerals and ranks highly in global production. Key minerals produced are coal, limestone, and iron ore. Major producing states are Jharkhand, Odisha, Chhattisgarh, Madhya Pradesh, Andhra Pradesh, and Karnataka. 2. Mining in India faces challenges of inadequate infrastructure, difficulties obtaining environmental clearances, and local community opposition. Several large projects have been delayed or withdrawn due to these issues. 3. Ore deposits form through various geological processes including magmatic, hydrothermal, metamorphic, and surficial processes. Theories of ore genesis explain the source of metals, how they are transported, and how they become trapped in

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1 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
2 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:
3 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
4 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
5 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.
6 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
7 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
8 Fig 3(a) Saddle reef Fig.3(b). Bendigo Goldfield, Victoria, Australia
9 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
10 ORE DEPOSITS and the Tectonic Cycle
11 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,
12 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.
13 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.
14 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) INFERRED SUB-ECONOMIC RESOURCES DEMONSTRATED SUB-ECONOMIC RESOURCES INFERRED MARGINAL RESERVES MARGINAL RESERVES INFERRED RESERVES RESERVES SPECULAT IVE HYPOTHETI CAL INDICATED MEASURED PROBABILITY RANGE INFERRED DEMONSTRATED UNDISCOVERED RESOURCES IDENTIFIED RESOURCES Economic Marginally Economic Sub- Economic Economic Feasibility Certainty Of Existence
15 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.
16  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
17 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.
18 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.
19 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
20 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 Adit
23 Source: Surface & UG excavations Ratan .R. Tatiya 2013
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 Sublevel Long hole Open stoping Cut-and-Fill Shrinkage VCR Sub Level Caving Block caving 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 massbehaviour. 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 ofproblemsofminenear-fieldstability 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 = 𝑃 𝑧𝑧(𝑊 𝑜 + 𝑊𝑃) Or 𝜎𝑝 = 𝑃 𝑧𝑧(𝑊 𝑜 + 𝑊 𝑝)/𝑊 𝑝 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
80 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 = 𝑊 𝑜/(𝑊 𝑜 + 𝑊𝑃) So that 1 − 𝑟 = 𝑊𝑃 𝑊 𝑜 + 𝑊𝑃 Insertion of this expression in the above equation, yields: 𝜎𝑃 = 𝑃 𝑧𝑧[ 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 = 𝑃 𝑧𝑧(𝑎 + 𝑐)(𝑏 + 𝑐) Or 𝜎𝑃 = 𝑃 𝑧𝑧(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 = 𝑃 𝑧𝑧[ 1 1 − 𝑟 ]
81 For a square pillar, of plan dimension WpxWp, are separated by rooms of dimension Wo, the equation is 𝜎𝑃 = 𝑃 𝑧𝑧[(𝑊 𝑜 + 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= 0.75 r= 0.9
82 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.
83 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) (%) (%) (%)
104 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
121 550 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
122 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
123  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
124 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
125 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.
126 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:
127 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.
128 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.
129 Fig. Breast stope (a). Schematic diagram of inclined room and pillar
130 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.
131 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.
132 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
133 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
134 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.
135 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.
136 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.
138 (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
139 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
142 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.
145 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.
146 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.
147 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.
148 (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.
149 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
150 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
151 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
152 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  Howard Hartman, Introductory Mining Engineering  B.H.G Bardy and E.T Brown., Rock Mechanics for underground mining  William Hustrulid and Richard L. Bullock Underground Mining Methods  Barton, N.R., Lien, R. and Lunde, J. 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mech. 6(4), 189-239.  Bieniawski, Z.T. 1989. Engineering rock mass classifications. New York: Wiley.  Brady, B.H.G. and Brown, E.T. 2006. Rock Mechanics for underground mining, 3rd Ed. The Netherlands: Springer.  Brown, E. T. 2003. Block Caving Geomechanics. Julius Kruttschnitt Mineral Research Centre: Brisbane.  Bullock, R. and Hustrulid, W. 2001. Chapter 3: Planning the Underground Mine on the Basis of Mining Method. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 29-48. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Hamrin, H. 2001. Chapter 1: Underground mining methods and applications. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 3–14. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Herne, V. and McGuire, T. 2001. Chapter 13: Mississippi Potash, Inc.’s, underground operations. In: Underground Mining Methods: Engineering Fundamentals and
153 International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 137-141. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Hoek, E. and Brown, E.T. 1997. Practical estimates of rock mass strength. Int. J. Rock Mech. Min. Sci. 34:8,1165-8,1186.  Krauland, N., Marklund, P.-I., and Board, M. 2001. Chapter 37: Rock support in cut-and- fill mining at the Kristineberg Mine. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 325-332. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Laubscher, D.H. 1990. A Geomechanics Classification System for the Rating of Rock Mass in Mine design. Journal of the South African Institute of Mining & Metallurgy, vol. 90, no 10, pp. 257-273.  Laubscher, D.H. 1994. Cave mining – the state of the art. The Journal of The South African Institute of Mining and Metallurgy, 94(10): 279-93.  Marchand, R., Godin, P., and Doucet, C. 2001. Chapter 19: Shrinkage stoping at the Mouska Mine. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 189-194. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Villaescusa, E. 1996. Excavation design for bench stoping at Mount Isa mine, Queensland, Australia. Trans. Instn Min. Metall., 105: A1–10.

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DMD LECTURE NOTES.pdf

  • 1. 1 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
  • 2. 2 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:
  • 3. 3 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
  • 4. 4 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
  • 5. 5 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.
  • 6. 6 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
  • 7. 7 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
  • 8. 8 Fig 3(a) Saddle reef Fig.3(b). Bendigo Goldfield, Victoria, Australia
  • 9. 9 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
  • 10. 10 ORE DEPOSITS and the Tectonic Cycle
  • 11. 11 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,
  • 12. 12 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.
  • 13. 13 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.
  • 14. 14 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) INFERRED SUB-ECONOMIC RESOURCES DEMONSTRATED SUB-ECONOMIC RESOURCES INFERRED MARGINAL RESERVES MARGINAL RESERVES INFERRED RESERVES RESERVES SPECULAT IVE HYPOTHETI CAL INDICATED MEASURED PROBABILITY RANGE INFERRED DEMONSTRATED UNDISCOVERED RESOURCES IDENTIFIED RESOURCES Economic Marginally Economic Sub- Economic Economic Feasibility Certainty Of Existence
  • 15. 15 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.
  • 16. 16  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
  • 17. 17 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.
  • 18. 18 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.
  • 19. 19 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
  • 20. 20 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. 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. 22 Fig . A-E different modes of access to deposits Fig. Cross-section of a service shaft Adit
  • 24. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 57
  • 58. 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 Sublevel Long hole Open stoping Cut-and-Fill Shrinkage VCR Sub Level Caving Block caving Magnitude of displacement in country rock Strain Energy storage in near-field rock Rock mass response to Mining
  • 59. 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 massbehaviour. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 70 Fig Plan view of a Room-and-Pillar stope Fig. Samsung limestone mines – South Korea
  • 71. 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. 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 ofproblemsofminenear-fieldstability 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. 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. 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. 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. 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. 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. 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. 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 = 𝑃 𝑧𝑧(𝑊 𝑜 + 𝑊𝑃) Or 𝜎𝑝 = 𝑃 𝑧𝑧(𝑊 𝑜 + 𝑊 𝑝)/𝑊 𝑝 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
  • 80. 80 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 = 𝑊 𝑜/(𝑊 𝑜 + 𝑊𝑃) So that 1 − 𝑟 = 𝑊𝑃 𝑊 𝑜 + 𝑊𝑃 Insertion of this expression in the above equation, yields: 𝜎𝑃 = 𝑃 𝑧𝑧[ 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 = 𝑃 𝑧𝑧(𝑎 + 𝑐)(𝑏 + 𝑐) Or 𝜎𝑃 = 𝑃 𝑧𝑧(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 = 𝑃 𝑧𝑧[ 1 1 − 𝑟 ]
  • 81. 81 For a square pillar, of plan dimension WpxWp, are separated by rooms of dimension Wo, the equation is 𝜎𝑃 = 𝑃 𝑧𝑧[(𝑊 𝑜 + 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= 0.75 r= 0.9
  • 82. 82 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.
  • 83. 83 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. 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. 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. 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. 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. 88 Fig. Shrinkage stope layout for Load-Haul-Dump (LHD) operations
  • 89. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 99 Figure Resign bolts (Brady and Brown) Figure : Grouted dowel in a grout –filled hole
  • 100. 100 Figure : Fully grouted cable bolt Figure Summary of development of different cable bolting systems in underground metal mines (Windsor, 1992, Brady & Brown)
  • 101. 101 Figure Summary of development of different cable bolting systems in underground metal mines (Windsor, 1992, Brady & Brown)
  • 102. 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. 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) (%) (%) (%)
  • 104. 104 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. 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. 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. 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. 108 Figure Vertical Crater Retreat method
  • 109. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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. 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
  • 121. 121 550 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
  • 122. 122 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
  • 123. 123  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
  • 124. 124 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
  • 125. 125 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.
  • 126. 126 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:
  • 127. 127 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.
  • 128. 128 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.
  • 129. 129 Fig. Breast stope (a). Schematic diagram of inclined room and pillar
  • 130. 130 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.
  • 131. 131 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.
  • 132. 132 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
  • 133. 133 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
  • 134. 134 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.
  • 135. 135 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.
  • 136. 136 Figure Underground mining capital and operating costs (Thomas W. Camm,1989)
  • 137. 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.
  • 138. 138 (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
  • 139. 139 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. 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. 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
  • 142. 142 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. 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. 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.
  • 145. 145 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.
  • 146. 146 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.
  • 147. 147 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.
  • 148. 148 (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.
  • 149. 149 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
  • 150. 150 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
  • 151. 151 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
  • 152. 152 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  Howard Hartman, Introductory Mining Engineering  B.H.G Bardy and E.T Brown., Rock Mechanics for underground mining  William Hustrulid and Richard L. Bullock Underground Mining Methods  Barton, N.R., Lien, R. and Lunde, J. 1974. Engineering classification of rock masses for the design of tunnel support. Rock Mech. 6(4), 189-239.  Bieniawski, Z.T. 1989. Engineering rock mass classifications. New York: Wiley.  Brady, B.H.G. and Brown, E.T. 2006. Rock Mechanics for underground mining, 3rd Ed. The Netherlands: Springer.  Brown, E. T. 2003. Block Caving Geomechanics. Julius Kruttschnitt Mineral Research Centre: Brisbane.  Bullock, R. and Hustrulid, W. 2001. Chapter 3: Planning the Underground Mine on the Basis of Mining Method. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 29-48. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Hamrin, H. 2001. Chapter 1: Underground mining methods and applications. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 3–14. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Herne, V. and McGuire, T. 2001. Chapter 13: Mississippi Potash, Inc.’s, underground operations. In: Underground Mining Methods: Engineering Fundamentals and
  • 153. 153 International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 137-141. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Hoek, E. and Brown, E.T. 1997. Practical estimates of rock mass strength. Int. J. Rock Mech. Min. Sci. 34:8,1165-8,1186.  Krauland, N., Marklund, P.-I., and Board, M. 2001. Chapter 37: Rock support in cut-and- fill mining at the Kristineberg Mine. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 325-332. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Laubscher, D.H. 1990. A Geomechanics Classification System for the Rating of Rock Mass in Mine design. Journal of the South African Institute of Mining & Metallurgy, vol. 90, no 10, pp. 257-273.  Laubscher, D.H. 1994. Cave mining – the state of the art. The Journal of The South African Institute of Mining and Metallurgy, 94(10): 279-93.  Marchand, R., Godin, P., and Doucet, C. 2001. Chapter 19: Shrinkage stoping at the Mouska Mine. In: Underground Mining Methods: Engineering Fundamentals and International Case Studies (eds W. A. Hustrulid and R. L. Bullock), 189-194. Society for Mining, Metallurgy and Exploration: Littleton, Colorado.  Villaescusa, E. 1996. Excavation design for bench stoping at Mount Isa mine, Queensland, Australia. Trans. Instn Min. Metall., 105: A1–10.