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Evaluation and
exploitation of gold
ore deposits at el Sukari area in eastern
desert of Egypt
By
B.Sc. Mining Eng. Project
2010/2011
A Graduate Project
Submitted in Partial Fulfillment of the Requirements for the
B.Sc. In Mining Engineering.
Department of Mining Engineering
Faculty of Petroleum and Mining Engineering
Suez Canal University
2011
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Submitted By:
Marketing
Ragab eid said
Sameh karam Hassan
Ore evaluation
Mohamed Mahmoud Mahmoud Afify
Mahmoud Mohamed Ahmed Eliwa
Opening up
Ahmed Ali Mohamed Ali
Hossam Eldin Hassan Ali
Drilling & Blasting
Ahmed Gamal Ahmed Hekal
Mohamed Ibrahim Mostafa
Design of transport system
Ibrahim Gharieb Ibrahim Dawood
Ragab Abd Elaziz Ibrahim Afify
Processing of R.O.M
Amir Mohamed Ahmed
Ahmed Hassanin Ahmed
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Supervised By:
Prof. Dr. Ali Hemeda Gomaa.
Prof. Dr. Mohamed Abd El Tawab El Gendy.
Prof. Dr. Saeed Abd Allah Mohamed.
Prof. Dr. Montaser Sabbah El Dein El Salmawy.
Prof.Dr. Mostafa Abas Hamam
Prof.Dr. Mohamed Hussien Allam.
Prof.Dr.Ahmed sedik
Dr. Abd El Azeem Mahmoud Abd El Aal.
Dr. Khaled Fayez
Prof.Dr. Salah Sameeh
Eng. Abd el menaem Selem
Eng. Amr Fathy
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ACKNOWLEDGEMENT
We have the great pleasure to express our deep appreciation and
thanks to all staff members of mining and minerals engineering
department, Suez Canal University Especially for Prof. Dr./
Mohamed Abd El Tawab El Gendy, Prof. Dr. \Saeed Abd Allah
Mohamed and Dr. Abd El Azeem Mahmoud Abd El Aal for their
advice and helpful discussion.
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Table of Contents
Geography .......................................................................................... - 7 -
Geology .............................................................................................. - 7 -
Marketing ......................................................................................... - 16 -
Ore Reserve Estimation ................................................................... - 44 -
Opening up ....................................................................................... - 99 -
Drilling and balsting ........................................................................ - 99 -
Loading and transportation ............................................................ - 145 -
Ore dressing ................................................................................... - 157 -
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Geography
The historic Sukari gold mine is at about 24°57'N 34°43'E, about 20 km from the
Red Sea, and 10 km south of a sealed
road that runs from Edfu on the Nile
to Marsa Alam on the Red Sea.
There is a recently opened
international airport about 70 km
north of Marsa Alam. Five-star
tourist resorts are now springing up
along the Red Sea coast for many
kilometres, almost overnight.
Geology
Theories of gold genesis:
Gold is concentrated by various
natural geologic processes to form commercial deposits of two principal types: hard rock
lode (primary) deposits and placer (secondary) deposits.
There three theories of gold genesis :-
1. Lode deposits are the gold deposits which remain locked within their original
solid rock formations. These are the targets for the "hard rock" prospector
seeking gold at the site of its deposition which was formed from mineralizing
solutions within the earth.
2. Another model which applies to the origins of some gold deposits, suggests
that gold-bearing solutions may be expelled from magma as it cools,
precipitating ore materials as they move into cooler surrounding rocks.
3. A third theory of gold genesis is applied mainly to gold-bearing veins in
metamorphic rocks that occur in mountain belts at continental margins.
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Gold-bearing minerals
Calaverite AuTe2, 44.5% gold; sylvanite (―graphic tellurium‖), AuTe3; native gold
alloyed with silver, etc
Gold most commonly occurs in quartz veins both as native, and as scales and wires
mechanically mixed in pyrite. It may be set free by oxidation and removal of pyrite.
It also accompanies mispickel, chalcopyrite, and rarely galena. On oxidizing and losing
tellurium, they yield extremely fine particles, not readily panned and resisting
amalgamation: called ―rusty‖ gold.
Predominant gold
A. Fissure veins containing native gold, alone, or mechanically mixed in pyrite and
much rarer base-metal sulphides, in quartz gangue. Gray, greasy-looking quartz seems
to accompany best values. Veins appear most frequently in schists, slates, or other
metamorphic rocks, and in association with intrusive rocks, of which granite is
commonest.
B. Impregnations and replacements of open-textured rocks with gold-bearing pyrite.
The ‖banket‖ of gold-bearing conglomerates of Transvaal (in South Africa) is the
best example.
C. Saddle reefs or arch-like deposits of gold-bearing quartz at crests of anticlines.
Saddle-reefs may succeed one another in depth. Slates or slaty schists are common
wall-rocks.
D. Veins carrying gold tellurides. We are associated with an eroded Eocene volcano,
often favoring neighborhood of minor dikes of phonolite and basaltic rocks, with which
volcanic activity closed. Purple fluorite is a characteristics associate.
E. Lateral impregnations and replacements of calcareous shales, with tellurides
along supply fissures, called verticals.
F. Contact zones, on the border of intrusive igneous rock and limestone, containing
gold-bearing mispickel in lime silicates. The usual contact zone of this type carries
copper sulphides with a little gold.
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G. Placer deposits of gold-bearing gravels, which may be: residual, from weathering
of rocks in situ; alluvial fans; sea-beaches with active surf; sea-beaches now elevated
and inland. Gold in streams favors places where current has been checked, as the inside
of bends; junctions of tributaries; heads of quiet reaches.
Sukari geology
The vein-type deposit is hosted in Late Neoproterozoic granite that intruded
island-arc and ophiolite rock assemblages. The vein-forming process is related to overall
late Pan-African shear and extension tectonics. At Sukari, bulk NE– SW strike-slip
deformation was accommodated by a local flower structure and extensional faults with
veins that formed initially at conditions of about 300 ºC and 1.5–2 kbar. Gold is
associated with sulfides in quartz veins and in alteration zones. Pyrite and arsenopyrite
dominate the sulfide ore beside minor sphalerite, chalcopyrite and galena. Gold occurs in
three distinct positions: (1) anhedral grains (GI) at the contact between As-rich zones
within the arsenian pyrite; (2) randomly distributed anhedral grains (GII) and along
cracks in arsenian pyrite and arsenopyrite, and (3) large gold grains (GIII) interstitial to
fine-grained pyrite and arsenopyrite. Fluid inclusion studies yield minimum
veinformation temperatures and pressures between 96 and 188ºC, 210 and 1,890 bar,
respectively, which is in the range of epi- to mesothermal hydrothermal ore deposits.
The structural evolution of the area suggests a longterm, cyclic process of repeated
veining and leaching followed by sealing, initiated by the intrusion of granodiorite.
General geology of the central Eastern Desert
The Neoproterozoic crust in NE Africa was consolidated by accretion of
intra-oceanic island arcs, continental micro-plates and oceanic plateaus. As a result of
convergent tectonics, a nappe assembly with two major tectonostratigraphic units was
established within the central Eastern Desert of Egypt: The structural basement, referred
to as ‗‗infrastructure’’ consists of orthogneisses, psammitic schists and amphibolites that
suffered amphibolite grade, polyphase metamorphic conditions. Structural cover units
summarized here as Pan-African Nappes and referred to as suprastructure‘‘ include
ophiolites, mélange lange-like sediments of accretionary-wedge type and calc-alkaline
volcanics similar to igneous rocks at modern arcs.
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Metamorphism in these units is of greenschist metamorphic grade. Exhumation of
previously buried high-grade structural basement units was achieved by combined
sinistral strike-slip faults and related north- and south-dipping, NE-trending normal
faults. This orogen-scale fault system is known as the Najd Fault System that
accommodated bulk NW–SE extension in the Arabian Nubian Shield. Overall NW–SE
extension exposed ‗‗gneissic domes‘‘ namely the Meatiq-Sibai- and Hafafit Domes (Fig.
1).
"Fig. 1 General geology and location of gold deposits of the central Eastern Desert. Note concentration of
metal occurrences close to major faults. Location of the Najd Fault System (NFS) is marked in the inset"
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Some of these domes have been interpreted as core complexes. Simultaneously with
exhumation of basement rocks, intramontane molasse basins with sediments delivered
from basement domes and PanAfrican Nappes were deposited. In addition, various
syn-tectonic and post-tectonic granitoids have been emplaced during late Pan-African
extension that softened the crust by enhanced advective heat supply.
Geology of the Sukari gold mine area
The mine occurs within a Late Neoproterozoic that intruded older
volcanosedimentary successions and an ophiolitic assemblage, both known as Wadi
Ghadir mélange. The volcanosedimentary succession is composed of andesites, dacites,
rhyodacites, tuffs and pyroclastics. Magmatic rocks are of calc-alkaline affinity and
were formed in an island-arc setting. The dismembered ophiolitic succession is
represented by a serpentinite at the base, followed upwards by a metagabbro-diorite
complex and sheeted dykes. Metagabbro-diorite rocks and serpentinites form lenticular
bodies (1–3 km2) as well as small bodies occur conformably scattered in the
volcanosedimentary arc assemblage. All rocks are weakly metamorphosed (lower
greenschist metamorphic facies), intensely sheared and transformed into various schists
along shear zones. Mineralized quartz veins and talc-carbonate veinlets are common.
The Sukari granitoid is elongated in a NNE direction and bounded from west and east by
two steep shear zones (Fig. 2B, C), covering an area of ca. 10 km. The fresh rock is
leucocratic, coarse-grained and pink in color. It has a heterogeneous mineralogical
composition and ranges from monzogranite to granodiorite with dominant quartz,
plagioclase and potash feldspars and less abundant biotite.
In the vicinity of shear zones the granite is foliated, elsewhere, however, it has sharp
intrusive contacts against the older rocks. Along those shear zones serpentinite and
andesite is altered to listvenite rock that attains up to 70 m in thickness and extends for
several kilometers. At the intersection of the two shear zones, where the gold
mineralization is concentrated, the Sukari granite is almost completely altered and
transected by a large amount of quartz veins (Fig. 2B).
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Fig. 2 A Simplified geology and structural frame of the
Sukari mine area. The system of NW-trending sinistral
strike-slip faults (lateral ramps and tear faults) and thrusts
(frontal ramps) forming the arc-shaped structure
developed during NW-directed tectonic thrust (black
arrow). The Sukari granite body intruded the system of
NE-trending sinistral shear zones with locally developed
flower structures. Inset: Flower structure block model of
the Sukari area looking northeast. Steep thrust faults,
strike-slip faults and orientation of major quartz veins
(qu) are indicated. A antithetic shear, T tension gash. B
Map of the Sukari mine area including shear zones and
quartz veins. C NW–SE section (1–2) across the Sukari
granite close to the mine. Shear zones are indicated
(modified from Khalaf and Oweiss 1993).
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Gold ore zone
Gold occurs in two textural positions and three generations in quartz veins or
veinlets: (1) as inclusions in pyrite and arsenopyrite (GI and GII) and (2) as interstitial
grains between pyrite and other
sulfides (GIII). Gold inclusions
(2–20 lm) in pyrite are either
located at the surfaces of As-rich
zones (GI) as revealed by BSE
images (Fig. 3A) or randomly
distributed. Gold inclusions in
arsenopyrite are randomly
distributed or located along
deformational cracks (GII, Fig.
3A). Interstitial gold grains
(GIII, Fig. 7B) are usually
associated with deformed pyrite
and arsenopyrite in the deformed and sheared smoky quartz (type Q2). In this textural
position, gold grains range from 2 to 80 lm and sometimes host small arsenopyrite and
pyrite crystals. Electron microprobe analysis (Table 1) revealed that gold is always
electrum (12–14 wt% silver).
No systematic compositional
difference between inclusion and
interstitial gold were detected.
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Structural controls to gold mineralization
The Sukari gold deposit is a large, sheeted vein-type and brittle-ductile shear zone
hosted gold deposit, developed in a late- to post-orogenic granitoid intrusive complex
intruded into the Neoproterozoic Hijaz Magmatic Arc of the Arabian-Nubian Shield.
Deformation at Sukari is manifest as a fold-thrust-nappe in the foreland to a large
metamorphic core complex - the Hafafit Culmination, uplifted at ca.680Ma and forming
part of the major Najd Fault System. Sukari lies on one of these arcuate, NE trending
thrusts. Gold mineralisation at Sukari (ca.530Ma) postdates the final stages of uplift, so
may not be related to Najd faulting event. Post-accretion shortening direction which
prevailed during the evolution of the Najd Fault System varied between NW-SE and
ESE-WNW.
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Marketing
History of gold
First smelting of gold
Egyptian goldsmiths carry out the first melting or fusing of ores in order
to separate the metals inside. They use blowpipes made from fire-resistant
clay to heat the smelting furnace.
2600 BC
Early gold jewellery
Goldsmiths of ancient Mesopotamia (modern-day Iraq) craft one of the
earliest pieces of gold jewellery, a burial headdress of lapis and carnelian
beads with willow leaf-shaped gold pendants.
1200-1500 BC
Advances in jewellery making
Artisans develop the lost-wax jewellery casting technique. The process
allows for improved hardness and colour variation which in turn broadens
the market for gold products.
1223 BC
Creation of Tutankhamen's funeral mask
Instantly recognised the world over, the funeral mask of Tutankhamun
is a triumph of gold craftsmanship from the ancient world.
600 BC
First gold dentistry practiced
The first use of gold in dentistry as the Etruscans begin securing
substitute teeth with gold wire. Bio-compatibility, malleability and
corrosion resistance still make gold valuable in dental applications.
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564 BC
First international gold currency created
King Croesus develops improved gold refining techniques, permitting
him to mint the world's first standardised gold currency. Their uniform gold
content allows 'Croesids' to become universally recognized and traded with
confidence.
300
First gold nanoparticles
The Romans use gold to colour the Lycurgus Cup. Melting gold powder
into glass diffuses gold nanoparticles throughout which then refract light,
giving the glass a luminous red glow.
1300
Hallmarking practice established
The world's first hallmarking system, scrutinising and guaranteeing the
quality of precious metal, is established at Goldsmith's Hall in London -
where London's Assay Office is still located today.
1370
The Great Bullion Famine begins
During the years 1370-1420, various major mines around Europe
become completely exhausted. Mining and production of gold declines
sharply throughout the region in a period known as 'The Great Bullion
Famine'.
1422
Venice's record year
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The Venice Mint strikes a record 1.2 million gold ducats using 4.26
metric tonnes of gold from Africa and Central Asia. These small coins
prove popular as they are easy to mint and carry plenty of value.
1511
Ferdinand unleashes invasion force
King Ferdinand of Spain proclaims "Get gold, humanely if you can, but
at all hazards, get gold!", launching unprecedented expeditions to the
Americas. Within years, the Inca and Aztec civilizations would be virtually
destroyed by Spanish conquerors.
1717
UK gold standard commences
Britain moves onto a de facto pure gold standard, as the government
links the currency to gold at a fixed rate (establishing a mint price of 77
shillings, ten and a half pennies per ounce of gold).
1803
First gold electroplating practiced
The first recorded experiment in electroplating is carried out by
Professor Luigi Brugnatelli at the University of Pavia. Gold electroplating
ensures improved conductivity, now essential to many 21st century
technologies.
1848
California Gold Rush begins
John Marshall discovers gold flakes while building a sawmill near
Sacramento, California. The greatest gold rush of all time follows as
40,000 diggers flock to California from around the World.
1885
South African Gold Rush begins
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While digging up stones to build a house, Australian miner George
Harrison finds gold ore on Langlaagte farm near Johannesburg. Miners
flock to the region. South Africa will go on to become the source of 40% of
the world's gold.
1885
First Faberge Easter egg crafted
Carl Faberge makes his first gold Imperial Easter Egg for Tsar
Alexander III. Named The Hen Egg, it was commissioned as a gift from the
Tsar to his wife, the Empress Maria Fedorovna, beginning a tradition that
lasts until 1917.
1870-1900
Adoption of gold standard
All major countries other than China switch to the gold standard, linking
their currencies to gold. The practice of bimetallism is abandoned.
1925
Gold standard returns
The UK returns to the gold standard at pre-war parity of $4.86=£1 with
sterling convertible to gold at 77sh 10.5d per standard ounce. This follows
the country's departure from the gold standard six years previously at the
outbreak of World War I.
1933
Roosevelt suspends gold
President Roosevelt suspends US dollar convertibility to gold (gold at
US$20.67/oz). The export of all transactions in, and the holding of gold by
private individuals, is forbidden. Presidential proclamation makes the
dollar convertible again in January 1934 at a new price of $35 per troy
ounce.
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1939
World War II closes gold market
The London gold market is closed on the outbreak of war, as at the
beginning of World War II. The world will later return to a fixed system of
exchange rates, this time with currencies fixed to the dollar and the dollar
convertible into gold.
1944
Bretton Woods conference
The Bretton Woods conference sets the basis of the post-war monetary
system. The US dollar is set to maintain a $35=1 oz gold conversion rate.
Other currencies are fixed in terms of US dollar, thus forming a Gold
Exchange Standard.
1961
First gold bonded microchips
Gold bonding wire is used in microchips engineered at Bell Labs in the
USA. Nowadays literally billions of chips are bonded this way every year,
controlling all manner of indispensible electrical devices.
1961
First gold in space
The first manned space flight uses gold to protect sensitive instruments
from radiation. In 1980, 41kgs of gold is included in space shuttle
construction through brazing alloys, fuel cell fabrication and electrical
contacts.
1967
First South African Krugerrand
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The Krugerrand is introduced in 1967, as a vehicle for private
ownership of gold. This iconic coin is actually intended for circulation as
currency.
1971
Gold window closed
The Bretton Woods system of fixed exchange rates comes to an end as
President Nixon "closes the gold window", suspending US dollar
convertibility to gold. The world enters its present day system of floating
exchange rates.
1985
First gold-based arthritis treatment
Pharmaceutical giant, SmithKline & French, develops Auranofin, a
gold-based drug for the treatment of rheumatoid arthritis. The drug receives
regulatory approval and goes on sale for the first time.
1999
First Central Bank Gold Agreement
The First Central Bank Gold Agreement (CBGA) is agreed. 15
European central banks declare that gold will remain an important element
of their reserves and collectively cap gold sales at 400 tonnes per year over
next five years.
2001
First gold used in heart surgery
Boston Scientific markets the first gold-plated stent (Niroyal) used in
heart surgery. Inserted inside large arteries and veins, such stents act like
scaffolding, propping open the blood vessels to allow adequate flow.
2003
K-gold launched in China
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The World Gold Council creates an entirely new market segment with
the launch of K-gold, the first 18k jewellery in China. The jewellery, in
predominantly white and yellow gold, takes its inspiration from Italian
design.
2004
Launch of SPDR® Gold Shares
The market is transformed by an innovative, secure and easy way to
access the gold market. Six years later SPDR® exceeds $55bn in assets
under management.
2009
Central banks return to buying
In the second quarter of the year, central banks collectively become net
purchasers of gold for the first time in two decades. This reflects a
combination of slowing sales from European banks and growing purchases
by emerging market countries.
2010
Gold price sustains record highs
Fiat currencies are undermined by inflation fears and successive
financial crises. The London pm fix achieves 35 separate successive highs
in the year to date.
2011
Gold in catalytic converters
Gold used in catalytic convertors by a leading European diesel car
manufacturer. The first use of gold in automotive emissions control
The Importance Of Gold In Egypt
One of the oldest elements of earth is the gold, which has been valued
since centuries. Especially the tribe, which found out gold, got many
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advantages out of the things gold brought with itself. This element might
not be as attractive as it is valued, but there are certain characteristics of
gold, which make it different from others.
Not only it is a harmless and comfortable element, gold tends to conduct
electricity quite well. With great convenience and comfort, it can be
transformed in to several shapes, and sizes. The history of the discovery of
gold is embedded in old books, and gives us an overview about it.
There is strong association between gold and Egypt, simply because this
country was one of the few civilizations to discover it. They benefited a
great deal from this metal, and it quickly began the resource of the country.
Not just in the olden days, but in the present day, gold holds the top most
value in the business hub.
It was a medium of exchange with aspect to price for purchasing and
selling commodities. Egypt was one of the countries that also with much
determination followed this concept. However, in Egypt, the role of gold
excelled beyond the boundaries of being used as money.
Some people are known to have worshipped the Sun. The people also
constitute the Egyptian civilization. Sun was often seen as a source of life.
For several civilisations, Gold was closely associated to the sun because of
its yellow and gleaming attributes. The Egyptians perceived gold as the
skin of Gods, especially the RA.
Besides the king, no one else was allowed to wear gold in those times.
Gradually and slowly, priests and royal members were also given the
privilege of wearing gold. The tomb of the king is known as the ―house of
gold‖, and the chamber is made out of gold.
Another amazing feature of gold was that it never rusts, which was
associated with the characteristics of god. The top of pyramids were
usually made out of a mixture of gold, and other metals. Due to its holy and
sacred value, it was often used to engrave the coffins. The mask of
Tutankhamun is an example of gold being used for funerary art. However,
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gold was not easy to dig out, and mining was quite a difficult task even
though there was rich gold present.
Due to the resource of gold present in the region, rivalries began
between Egypt and its bordering countries while mining for gold. The army
forces strictly monitored the convicts who were given the task of gold
mining. Nubia and the Eastern desert were abundant with gold; hence, all
the mining took place there. Looking at the current popularity of jewelers
in Egypt proves that the value and respect for gold is still alive.
Characteristics
• Streak : yellow
• Hardness : 2.5 -3
• Sp.gr :15.5 – 19.3
• Color : gold – yellow to brass yellow
• Luster : metallic
• Cleavage : none
• Fracture: soft
• soluble: in aquaregia and mercury
• crystals: cubic
• similar to : pyrite , chalcopyrite , biotite , markasite
• accompanied by: pyrite , chalcopyrite , sphalerite , magnetite ,
quartiz , tourmaline
Some Uses of gold
Jewelry:
Alloys with lower caratage, typically 22k, 18k, 14k or 10k, contain
higher percentages of copper, or other base metals or silver or palladium in
the alloy.
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Medicine:
Gold alloys are used in restorative dentistry, especially in tooth
restorations, such as crowns and permanent bridges.
Industry:
Gold solder is used for joining the components of gold jewelry by
high-temperature hard soldering or brazing.
Electronics:
The concentration of free electrons in gold metal is 5.90×1022 cm−3.
Gold is highly conductive to electricity, and has been used for electrical
wiring in some high-energy applications. gold has the advantage of
corrosion resistance
Commercial chemistry:
Gold is attacked by and dissolves in alkaline solutions of potassium or
sodium cyanide, to form the salt gold cyanide—a technique that has been
used in extracting metallic gold from ores in the cyanide process. Gold
cyanide is the electrolyte used in commercial electroplating of gold onto
base metals and electroforming.
Investing in Gold Mines:
Many holders of gold store it in form of bullion coins or bars as a
hedge against inflation or other economic disruptions. However, some
economists do not believe gold serves as a hedge against inflation or
currency depreciation
An alternative to buying physical gold bullion is to invest in the shares
of companies that are involved in the exploration and mining of
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gold. Analysts consider this to be a more risky investment, as the chances
of making a large return on your investment will depend both on the price
of gold price and the success of that company. However, in theory as the
price of gold bullion goes up, so does the value of gold mining shares.
As this is a highly technical area, if you are considering this kind of
investment we recommend doing plenty of research on the industry and
speaking to an experienced financial advisor. As with all investments,
financial advisors would encourage a diverse portfolio, so it is prudent to
consider a complimentary investment of gold bullion too. Overall, though
an investment in mining shares can be quite risky, it can provide a very
good return on investment.
World gold productive countries
Rank Country/Region Gold production (kilograms)
World 2,310,000
1 South Africa 272,128
2 China 247,200
3 Australia 247,000
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4 United States 242,000
5 Peru 203,268
6 Russia 159,340
7 Canada 104,198
8 Mali 85,411
9 Uzbekistan 84,000
10 Ghana 66,205
11 Indonesia 58,773
12 Papua New Guinea 58,349
13 Argentina 44,131
14 Chile 42,100
15 Brazil 40,075
16 Tanzania 39,750
17 Philippines 36,098
18 Mexico 35,899
19 Mongolia 22,561
20 Guinea 16,336
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Gold Production By Country (million of ounces)
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50% of all gold ever produced was produced since 1960
80% of all gold ever produced was produced since 1900
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Top Gold Importers
$10.9 billion (33.1% of top ten
gold importers, up 6% from 2004)
India
$4.4 billion (13.5%, up 10.9%) United
States
$3.9 billion (11.8%, up 11.4%) Turkey
$3.5 billion (10.5%, down 2.8%) Italy
$2.3 billion (6.9%, up 60.1%) Canada
$2.1 billion (6.5%, up 18.5%) Australia
$2 billion (6%, up 71.4%) Thailand
$1.4 billion (4.1%, up 9.2%) Malaysia
$1.3 billion (4.1%, up 12.1%) Japan
$1.2 billion (3.5%, up 17.9%). Germany
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Top Gold Exporters
In 2005,
the following nations exported the most gold by value.
United States $ 5.5 billion (23.2% of top ten gold
exporters, up 25.2% from 2004)
Australia $4.4 billion (18.6%, up 7.1%)
Canada $3.7 billion (15.5%, up 28.4%)
Peru $3.1 billion (12.9%, up 30.1%)
Hong Kong $2.8 billion (11.8%, down 55.9%)
Japan $1.4 billion (6%, up 17.8%)
Germany $841.7 million (3.5%, up 22.7%)
Singapore $764.8 million (3.2%, up 14.2%)
Italy $628.7 million (2.6%, up 41%)
Colombia $627.2 million (2.6%, up 9%).
Italy, Peru, America and Germany had the fastest increasing exports for
gold in 2005. However, while higher gold prices in 2008 may lead to
more demand for gold bullion and gold mining stock investments, gold
exports may well decrease as global trade partners wait for gold prices to
normalize.
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Prices in last 10 years
prices in last 5 years
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Prices last year
presence of gold in the eastern desert
we can find the gold deposits in many groups :
1-North group :
It includes 30 sites in between 27 45 and 28 00 latitude32 45 and 33 05
longitudes
2-middle group :
It includes 63 sites in between 24 10 and 26 45 latitude33 15 and 35 20
longitudes
3-southern east group :
It includes 7 sites in between 22 15 and 23 20 latitude And 34 25 and 35
50 longitudes
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4-southern west group :
It includes 19 sites in between 22 00 and 23 00 latitude
Some sites in details
1-Atoud
Location :
55 km south west of marsa alam
Intersection between 24 58 latitude with 34 40 longitudes
Reserve :
Proven : 8595 tons which have 109 kg gold
Possible : 10895 tons which have 78 kg gold
Probable : 13600 tons which have 238 kg
2-Barramiya
Location : 105 km east of edfu city
Intersection between 25 05 latitude and 33 47 longitudes
Reserve :
It contains 16 million tons ore which have 21 tons gold
3-Hangaliya
Location : 80 km south west of marsa alam
Reserve :
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478490 tons ore which have at least 478 kg gold
4- sukari
Location :
30 km south east of marsa alam
Reserve :
1.2 million tons ore which contain 2447 kg gold
5- umm ulaygah
Location :
80 km south west of ras bnas
Reserve :
Not calculated yet
6- Umm EL Rus
Location :
80 km south of qusair city
Reserve :
Not calculated yet
7-anait
Location :
160 km south of abo ghsoun
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Reserve :
83000 tons ore which contain 185 kg gold
8-Umm Ud
Location :
55 km south west of marsa alam
Reserve :
15600 tons ore which contains 354 kg gold
9-Hamash
Location :
120 km south west of marsa alam
the production from the beginning of the project until February 2010 is
65 kg and the plan of 2010 refers to the production of 15000 ounces
10 –Umm higab
Location :
40 km north of Hamash , 35 km south east of Barramiya
11-samut
45 km south east of Barramiya , 35 km south west of umm higab
12-Umm samra
60 km north east of Barramiya
Gold content 0.5 – 12 gm\ton
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13- abu marawat
Location :
North east abu marawat valley
Reserve :
290,000 tons ore which have 1210 kg gold
14-Hamama
Location :
In the road between Qena and Safaga
Reserve :
Not calculated yet
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. Gold deposits and occurrences in the Eastern Desert of Egypt (compiled
from Kochine and Basyuni, 1968). (1) Umm mongul; (2) Umm Balad; (3)
Wadi Dib; (4) Fatira; (5) Abu Marawat; (6) Wadi Gasus; (7) Semna; (8)
Gebel Semna; (9) Abu Qarahish; (10) Kab Amiri; (11) Sagi; (12) Gidami;
(13) Hamama; (14) Erediya; (15) Abu Had; (16) Atalla; (17) Rebshi; (18)
Umm Esh; (19) Fawakhir; (20) Hammamat; (21) Umm Had; (22) EL Sid;
(23) Umm Selimat; (24) Hammuda; (25) EL Nur; (26) Kareim; (27) Kab EL
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Abyad; (28) Tarfawi; (29) Sherm ELBahaari; (30) Zeidum; (31) Wadi
Zeidum; (32) Umm Rus; (33) Sigdit; (34) Talat Gadalla; (35) Abu
Muawaad; (36) Daghbag; (37) EL Hisinat; (38) Bokari; (39) Umm Samra;
(40) Abu Dabbab; (41) Abu Qaria; (42) Umm Saltit; (43) Bezah; (44) Umm
Selim; (45) Barramiya; (46) Dungash; (47) Samut; (48) Umm Hugab; (49)
Urf EL Fahid; (50) Atud; (51) Sukkari; (52) Umm Tundeba; (53)
Hanglaliya; (54) Kurdeman; (55) Sabahia; (56) Umm Ud; (57) Allawi; (58)
Lewewi; (59) Dweig; (60) Hamash; (61) Geli; (62) Qulan; (63) Kab EL
Rayan; (64) Sheialik; (65) AbuRahaya; (66) Wadi Khashb; (67) Umm
Eleiga; (68) Betan; (69) Qurga Rayan; (70) Hutit; (71) Kalib; (72) Kurtunos;
(73) EL Hudi; (74) Hariari; (75) Um Shira; (76) Neqib; (77) Haimur; (78)
The Nile Valley (Block E); (79) Umm Garaiart; (80) Marahib; (81) Atshani;
(82) Murra; (83) Filat; (84) Seiga I; (85) Seiga II; (86) Umm Shashoba; (87)
Abu Fass; (88) Umm Tuyur; (89) Betam; (90) Umm Egat; (91) Kurbiai; (92)
Romit.
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Ore Reserve Estimation
Introduction
Ore reserve estimates are assessments of the quantity and tenor of a mineral that
may be profitably and legally extracted from a mineral deposit through mining and/or
mineral beneficiation. (Examination and evaluation of ore deposites. Generally it
means the determination of the extent and value of mineral deposites.)
Practical consideration of Mineral Resources and Mineral Reserves
Mineral Resource confidence classification should take into account practical
considerations such as drilling, sampling and assay integrity, drill hole spacing,
geological control and continuity, grade continuity, estimation method and block size,
potential mining method and reporting period. Ore Reserve confidence classification
should take into account the confidence classification of the Mineral Resource and
should not include Inferred Resources. Cut-off grades, mining and metallurgical
factors or assumptions, cost and revenue factors, market assessment (where
appropriate) and other risk factors such as environmental, social or political should be
considered by the CP in terms of their impact on confidence in the Ore Reserve
estimate.
Mineral Resource
A concentration or occurrence of material of intrinsic economic interest in or on the
Earth‘s crust in such form, quality and quantity that there are reasonable prospects for
eventual economic extraction. The location, quantity, grade, geological characteristics
and continuity of a Mineral Resource are known, estimated or interpreted from specific
geological evidence and knowledge. Mineral Resources are sub-divided, in order of
increasing geological confidence, into Inferred, Indicated and Measured categories.
Measured Mineral Resource
That part of a Mineral Resource for which tonnage, densities, shape, physical
characteristics, grade and mineral content can be estimated with a high level of
confidence. It is based on detailed and reliable exploration, sampling and testing
- 45 - | P a g e
information gathered through appropriate techniques from locations such as outcrops,
trenches, pits, workings and drill holes. The locations are spaced closely enough to
confirm geological and grade continuity.
Ore Reserve
The economically mineable part of a Measured and/or Indicated Mineral
Resource. It includes diluting materials and allowances for losses, which may occur
when the material is mined. Appropriate assessments and studies have been carried out,
and include consideration of and modification by realistically assumed mining,
metallurgical, economic, marketing, legal, environmental, social and governmental
factors. These assessments demonstrate at the time of reporting that extraction could
reasonably be justified. Ore Reserves are sub-divided in order of increasing confidence
into Probable Ore Reserves and Proved Ore Reserves.
Probable Ore Reserve
The economically mineable part of an Indicated, and in some circumstances, a
Measured Mineral Resource. It includes diluting materials and allowances for losses
which may occur when the material is mined. Appropriate assessments and studies
have been carried out, and include consideration of and modification by realistically
assumed mining, metallurgical, economic, marketing, legal, environmental, social and
governmental factors. These assessments demonstrate at the time of reporting that
extraction could reasonably be justified.
Proved Ore Reserve
The economically mineable part of a Measured Mineral Resource. It includes
diluting materials and allowances for losses which may occur when the material is
mined. Appropriate assessments and studies have been carried out, and include
consideration of and modification by realistically assumed mining, metallurgical,
economic, marketing, legal, environmental, social and governmental factors. These
assessments demonstrate at the time of reporting that extraction could reasonably be
justified.
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Mineral Resources and Mineral Reserves must be reported on a site by site basis.
Where estimates for both Mineral Resources and Mineral Reserves are reported, for
consistency, a single form of reporting should be used in a report. Appropriate forms
of clarifying statements may be:
The Measured and Indicated Mineral Resources are inclusive of those
Mineral Resources modified to produce the Mineral Reserves,‘ or
The Measured and Indicated Mineral Resources are additional to the Mineral
Reserves.‘
Inferred Mineral Resources are, by definition, always additional to Mineral
Reserves.
RESOURCE ESTIMATION METHODOLOGY
A resource estimate is based on prediction of the physical characteristics of a mineral
deposit through collection of data, analysis of the data, and modeling the size, shape,
and grade of the deposit. Important physical characteristics of the ore body that must be
predicted include:
(1) The size, shape, and continuity of ore zones,
(2) The frequency distribution of mineral grade,
(3) The spatial variability of mineral grade.
These physical characteristics of the mineral deposit are never completely known, but
are inferred from sample data.
The sample data consist of one or more of the following
1. Physical samples taken by drilling, trenching, test pitting, and channel sampling.
2. Measurement of the quantity of mineral in the samples through assaying or other
procedures.
3. Direct observations such as geologic mapping and drill core logging.
- 47 - | P a g e
Estimation of the resource requires analysis and synthesis of these data to develop a
resource model.
Methods used to develop the resource model may include
1. Compilation of the geologic and assay data into maps, reports, and computer
databases.
2. Delineation of the physical limits of the deposit based on geologic interpretation of
the mineralization controls at a reasonable range of mining cutoff grades.
3. Compositing of samples into larger units such as mining bench height, seam
thickness, or minable vein width.
4. Modeling of the grade distribution based on histograms and cumulative frequency
plots of grades.
5. Evaluation of the spatial variability of grade using experimental variograms.
6. Selection of a resource estimation method and estimation of quantity and grade of
the mineral resource.
The estimation procedure must be made with at least minimal knowledge of the
proposed mining method since different mining methods may affect the size, shape,
and/or grade of the potentially minable ore reserve. The most important mining
factors for consideration in evaluation of the ore reserve from the resource are:
The range of likely cutoff grades. The degree of selectivity and the size of the
selective mining unit for likely mining methods. Variations in the deposit that affect the
ability to mine and/or process the ore.
These mining factors often determine the degree of detail that is required for the
resource model and thus the degree of difficulty to develop a resource model for
estimating ore reserves. For example, a disseminated gold deposit may be continuous
and regular in shape, if mined by bulk, open pit methods. The same deposit may be
discontinuous and difficult to estimate, however, if mined by more selective
underground methods at a higher cutoff grade. Such large differences in deposit shape
due to variations in cutoff grade and mining method may require different ore reserve
estimation methods or different mining methods.
- 48 - | P a g e
DATA COLLECTION AND GEOLOGIC INTERPRETATION
Data that must be collected and compiled for the resource estimate are as follows:
1. Reliable assays from an adequate number of representative samples.
2. Coordinate locations for the sample data.
3. Consistently recorded geologic data that describe the mineralization controls.
4. Cross sections or plan maps with the geologic interpretation of the mineralization
controls.
5. Tonnage factors or specific gravities for the various ore and waste rock categories.
6. Surface topographic map, especially for deposits to be surface mined.
Although small deposits may be evaluated manually using data on maps and in reports,
the amount of data required for a resource estimate is often large, and data may be more
efficiently evaluated if they are entered into a computer database. Computer programs
can then be used to retrieve the data for printing reports, plotting on digital plotters,
statistical analysis, and resource estimation. Minimum information that should be in-
cluded in a drillhole database are :
1. Drillhole number or other identification.
2. Hole length, collar coordinates, and down-hole surveys.
3. Sample intervals and assay data.
4. Geologic data such as lithology, alteration, oxidation, etc.
5. Geotechnical data such as RQD (rock quality designation).
Types of reserve calculation
The classical, employing two-dimensional maps and hand calculation. The
geostistical, a more soghisticaded approach requiring digital computers to prepare
statistically derived estimates. Different classical procedures are used in calculating
reserve estimates. They differ mainly in the ways in which they combine the sample
data.
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Reserve estimation from maps methods
1. polygon method
2. The triangle method
3. In the section method
polygon method
In the section method
- 51 - | P a g e
The triangle method
Types of manual methods
1. Vertical sections
2. Horizontal sections
3. Block module
The method used to calculation
Vertical section:
Av. Assay for each bore hole:
Av. Au = ∑ (D x Au) / ∑ D
Av. Assay for every section:
Av. Au = ∑ (W x D x Au)/ ∑ (W x D)
Procedures:
The procedure for the calculation of a reserve estimate – tonnage and average
grade of ore for a mineral deposit is demonstrated in the case study.
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Sec. 10000 mN
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1321 10000 10965 3 1.11 3.33 1.08 35 105 116.55
sum
9 1.07 9.63 35 315 337.05
12 12.96 420 453.6
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD99 10000 371.3 40 2.86 114.4 3.82104167 35 1400 4004
1 12.2 12.2 35 35 427
6 8.09 48.54 35 210 1698.9
1 8.27 8.27 35 35 289.45
sum 48 183.41 1680 6419.35
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD416 10000 10523 23 6.92 159.16 8.12410256 35 805 5570.6
sum
11 12.53 137.83 35 385 4824.05
5 3.97 19.85 35 175 694.75
39 316.84 1365 11089.4
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1173 10000 10675 26 3.59 93.34 4.11560606 35 910 3266.9
sum
1 27.3 27.3 35 35 955.5
1 35.7 35.7 35 35 1249.5
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33 2.37 78.21 35 1155 2737.35
1 8.75 8.75 35 35 306.25
1 9.32 9.32 35 35 326.2
2 6.27 12.54 35 70 438.9
1 6.47 6.47 35 35 226.45
66 271.63 2310 9507.05
sec 10000N Sum 5775 27469.4
average assay 4.756606061
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Sec. 10050 mN
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1313 10050 10665 45 1.07 48.15 1.07 50 2250 2407.5
sum 45 48.15 2250 2407.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1388 10050 10550 3 1.54 4.62 9.96142857 50 150 231
sum
20 7.39 147.8 50 1000 7390
5 25.3 126.5 50 250 6325
28 278.92 1400 13946
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD118
9 10050 10560 20 2.46 49.2 2.46 50 1000 2460
sum 20 49.2 1000 2460
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD126
3 10050 10625 2 4.1 8.2 3.88 50 100 410
sum
65 3.96 257.4 50 3250 12870
3 2 6 50 150 300
70 271.6 3500 13580
sec 10050N Sum 8150 32393.5
average assay 3.974662577
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Sec. 10100 mN
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1328 10100 10660 19 8.03 152.6 4.760714 50 950 7628.5
sum
7 20.58 144.1 50 350 7203
72 2.36 169.9 50 3600 8496
98 466.6 4900 23328
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1166 10100 10611 23 6.23 143.3 2.986923 50 1150 7164.5
68 1.89 128.5 50 3400 6426
91 271.8 4550 13591
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1174 10100 10665 104 1.59 165.4 1.713396 50 5200 8268
sum
2 8.13 16.26 50 100 813
106 181.6 5300 9081
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1225 10100 10662 83 1.95 161.9 2.612667 50 4150 8092.5
sum
1 5.62 5.62 50 50 281
3 15.83 47.49 50 150 2374.5
2 5.75 11.5 50 100 575
1 8.68 8.68 50 50 434
90 235.1 4500 11757
- 57 - | P a g e
sec 10100 Sum 19250 57756
average assay 3.000311688
- 58 - | P a g e
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Sec. 10150 mN
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1289 10150 10725 4 1.62 6.48 2.045 50 200 324
Sum
13 2.17 28.21 50 650 1410.5
1 8 8 50 50 400
8 1.31 10.48 50 400 524
26 53.17 1300 2658.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD500 10150 10556 56 2.03 113.68 2.03 50 2800 5684
Sum 56 113.68 2800 5684
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1165 10150 10662 45 2.35 105.75 2.35 50 2250 5287.5
Sum 45 105.75 2250 5287.5
sec 10150N Sum 6350 13630
average assay 2.146456693
- 61 - | P a g e
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Sec. 10200 mN
Hole N E D(m) Au
(g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1383 1020
0
1068
5 3 2.82 8.46 2.82 50 150 423
Sum 3 8.46 150 423
Hole N E D(m) Au
(g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1880 1020
0
1069
0 17 29.7 504.9 51.9421053 50 850 25245
sum
2 241 482 50 100 24100
19 986.9 950 49345
Hole N E D(m) Au
(g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RD502 1020
0
1055
7 38 6.86
260.6
8 9.2288 50 1900 13034
sum
9 19.6
1
176.4
9 50 450 8824.5
3 8.09 24.27 50 150 1213.5
50 461.4
4 2500 23072
Hole N E D(m) Au
(g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD11
61
1020
0
1064
9 23 2.6 59.8 3.76884615 50 1150 2990
- 62 - | P a g e
sum
3 12.7
3 38.19 50 150 1909.5
26 97.99 1300 4899.5
Hole N E D(m) Au
(g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD11
77
1020
0
1070
6 65
18.3
1
1190.
2 31.9614286 50 3250 59507.5
sum
5 209.
4
1047.
2 50 250 52357.5
70 2237.
3 3500 111865
sec 10200N Sum 8400 189605
average assay 22.57196429
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Sec. 10250 mN
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1351 10250 10678 34.4 6 206.4 6.3673165 50 1720 10320
Sum
1.6 85.88 137.408 50 80 6870.4
2 15.95 31.9 50 100 1595
11 3.91 43.01 50 550 2150.5
2 14.4 28.8 50 100 1440
1 9.98 9.98 50 50 499
26 1.51 39.26 50 1300 1963
1 6.26 6.26 50 50 313
79 503.018 3950 25150.9
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1386 10250 10595 37 1.04 38.48 8.3125333 50 1850 1924
sum
5 1.08 5.4
50 250 270
7 1.25 8.75 50 350 437.5
3 6.03 18.09 50 150 904.5
22 14.26 313.72 50 1100 15686
1 239 239 50 50 11950
75 623.44 3750 31172
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
DGT408 10250 10429 26 4.23 109.98 4.8957143 50 1300 5499
sum 2 13.55 27.1 50 100 1355
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28 137.08 1400 6854
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1162 10250 10700 61 2.27 138.47 2.9880882 50 3050 6923.5
sum
2 5.48 10.96 50 100 548
1 13.6 13.6 50 50 680
4 10.04 40.16 50 200 2008
68 203.19 3400 10159.5
sec 10250N sum 12500 73336.4
avg assay 5.866912
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Sec. 10300 mN
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D340 10300 10620 19 5.85 111.15 5.85 50 950 5557.5
Sum 19 111.15 950 5557.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1304 10300 10831 15 2.88 43.2 3.3413793 50 750 2160
sum
1 25 25
50 50 1250
12 1.74 20.88 50 600 1044
1 7.82 7.82 50 50 391
29 96.9 1450 4845
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RC415 10300 10445 10 3.64 36.4 2.96 50 500 1820
sum
17 2.56 43.52 50 850 2176
27 79.92 1350 3996
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD388 10300 10644 30 3.65 109.5 11.994925 50 1500 5475
sum
3 20.8 62.34
50 150 3117
24 9.7 232.8 50 1200 11640
1 185 185 50 50 9250
9 23.8 214.02 50 450 10701
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67 803.66 3350 40183
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1125 10300 10700 49 1.86 91.14 2.4228 50 2450 4557
sum
1 30 30 50 50 1500
50 121.14 2500 6057
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1178 10300 10758 105 4.1 430.5 5.4367033 50 5250 21525
sum
17 12.5 211.65
50 850 10583
2 76.4 152.7 50 100 7635
3 7.49 22.47 50 150 1123.5
1 14 14 50 50 700
51 2.04 104.04 50 2550 5202
1 30.6 30.6 50 50 1530
2 11.8 23.52 50 100 1176
182 989.48 9100 49474
sec 10300N sum 18700 110113
average assay 5.888368984
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- 71 - | P a g e
Sec. 10350 mN
Hole N E D(m) Au
(g/t)
AU *
D
Avg hole
AU.
DIST.BET
holes w*D w*D*AU
D1284 10350 10753 11 2.98 32.78 2.0438554 50 550 1639
sum
1 6.71 6.71
50 50 335.5
1 6.72 6.72 50 50 336
5 1.9 9.5 50 250 475
35 1.58 55.3 50 1750 2765
1 6.55 6.55 50 50 327.5
1 9.09 9.09 50 50 454.5
26 1.13 29.38 50 1300 1469
1 5.03 5.03 50 50 251.5
1 8.58 8.58 50 50 429
83 169.64 4150 8482
Hole N E D(m) Au
(g/t)
AU *
D
Avg hole
AU.
DIST.BET
holes w*D w*D*AU
D1345 10350 10730 94 3.98 374.12 3.3041036 50 4700 18706
sum
3 27.84 83.52
50 150 4176
1 5.18 5.18 50 50 259
4 11.56 46.24 50 200 2312
1 38 38 50 50 1900
1 12.5 12.5 50 50 625
7 10.52 73.64 50 350 3682
54 1.64 88.56 50 2700 4428
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4 9.62 38.48 50 200 1924
3 2.13 6.39 50 150 319.5
43 0.87 37.41 50 2150 1870.5
15 0.72 10.8 50 750 540
21 0.69 14.49 50 1050 724.5
251 829.33 12550 41466.5
Hole N E D(m) Au
(g/t)
AU *
D
Avg hole
AU.
DIST.BET
holes w*D w*D*AU
D1366 10350 10640 58 2.53 146.74 10.348934 50 2900 7337
sum
2 23.34 46.68
50 100 2334
1 27.4 27.4 50 50 1370
1 6.86 6.86 50 50 343
2 6.72 13.44 50 100 672
10 45.94 459.4 50 500 22970
2 227.7 455.4 50 100 22770
39 1.76 68.64 50 1950 3432
7 5.43 38.01 50 350 1900.5
122 1262.6 6100 63128.5
Hole N E D(m) Au
(g/t)
AU *
D
Avg hole
AU.
DIST.BET
holes w*D w*D*AU
RCD112
4 10350 10700 42 6.24 262.08 8.9569444 50 2100 13104
sum 1 213 213 50 50 10650
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27 3.66 98.82 50 1350 4941
1 26 26 50 50 1300
1 45 45 50 50 2250
72 644.9 3600 32245
sec 10350 sum 26400 145322
average assay 5.504621212
- 73 - | P a g e
- 74 - | P a g e
Section10350
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1143 10400 10710 212 2.37 502.44 3.3339631 50 10600 25122
sum
1 8 8
50 50 400
1 94.2 94.2 50 50 4710
3 39.61 118.83 50 150 5941.5
217 723.47 10850 36173.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1307 10400 10845 19 5.31 100.89 8.315 50 950 5044.5
2 9.27 18.54 50 100 927
1 63.5 63.5 50 50 3175
sum 22 182.93 1100 9146.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD523 10400 10696 6 5.09 30.54 5.9487952 50 300 1527
1 15.8 15.8
50 50 790
9 4.07 36.63 50 450 1831.5
0.6 26.3 15.78 50 30 789
sm 16.6 98.75 830 4937.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
- 75 - | P a g e
RCD1186 10400 10749 36 1.91 68.76 2.77625 50 1800 3438
sum
1 5.32 5.32
50 50 266
1 5.87 5.87 50 50 293.5
1 6.3 6.3 50 50 315
1 16.1 16.1 50 50 805
16 3.32 53.12 50 800 2656
56 155.47 2800 7773.5
sec 10400N sum 15580 58031
average assay 3.724711168
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D338A 10450 10506 5 14.76 73.8 14.76 50 250 3690
sum 5 73.8 250 3690
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1073 10450 10751 197 2.18 429.46 2.8535149 50 9850 21473
sum
3 12.35 37.05
50 150 1852.5
2 54.95 109.9 50 100 5495
202 576.41 10100 28820.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD508 10450 10636 60 2.4 144 2.4 50 3000 7200
sum 60 144 3000 7200
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1187 10450 10750 51 4.45 226.95 6.6868421 50 2550 11347.5
sum
1 36 36
50 50 1800
9 15.51 139.59 50 450 6979.5
14 4.44 62.16 50 700 3108
1 43.5 43.5 50 50 2175
76 508.2 3800 25410
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1187 10450 10742 4 5.84 23.36 1.846405 50 200 1168
sum
1 20 20
50 50 1000
2 10.52 21.04 50 100 1052
3 1.41 4.23 50 150 211.5
38.4 0.54 20.736 50 1920 1036.8
48.4 89.366 2420 4468.3
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1235 10540 10800 8 4.7 37.6 7.0777778 50 400 1880
sum
1 26.1 26.1 50 50 1305
9 63.7 450 3185
sec 10450N sum 20020 72773.8
average assay 3.635054945
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1042 10500 10767 75 2.17 162.8 2.784937 50 3750 8138
sum
1 7.37 7.37
50 50 368.5
3 16.63 49.89 50 150 2495
79 220 3950 11001
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1231 10500 10795 4 2.93 11.72 2.010952 50 200 586
sum
4 3.14 12.56
50 200 628
34 1.77 60.18 50 1700 3009
42 84.46 2100 4223
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1320 10500 10875 2 2.44 4.88 6.288 50 100 244
sum
2 11.24 22.48
50 100 1124
1 4.08 4.08 50 50 204
5 31.44 250 1572
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD509 10500 10608 34 3.08 104.7 3.08 50 1700 5236
sum 34 104.7 1700 5236
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
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RCD1213 10500 10500 28 2.56 71.68 4.865902 50 1400 3584
sum
3 10.75 32.25
50 150 1613
26 4.28 111.3 50 1300 5564
1 18.7 18.7 50 50 935
3 20.97 62.91 50 150 3146
61 296.8 3050 14841
sec 10500N sum 11050 36873
average assay 3.336877828
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1013 10550 10700 43 1.61 69.23 3.17228571 50 2150 3461.5
sum
1 6.81 6.81
50 50 340.5
22 3.83 84.26 50 1100 4213
2 11.73 23.46 50 100 1173
2 19.15 38.3 50 100 1915
70 222.06 3500 11103
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1280 10550 10843 24 0.93 22.32 99.4226087 50 1200 1116
sum
46 2.21 101.66
50 2300 5083
2 6.47 12.94 50 100 647
1 5.71 5.71 50 50 285.5
1 11.3 11.3 50 50 565
3 5.74 17.22 50 150 861
35 164.09 5743.2 50 1750 287158
1 5420 5420 50 50 271000
2 49.65 99.3 50 100 4965
115 11434 5750 571680
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1280 10550 10843 24 0.93 22.32 1.6787931 50 1200 1116
sum 31 1.82 56.42 50 1550 2821
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2 6.46 12.92 50 100 646
1 5.71 5.71 50 50 285.5
58 97.37 2900 4868.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1379 10550 10770 4 1.66 6.64 1.86862605 50 200 332
sum
3 1.17 3.51
50 150 175.5
2 1.85 3.7 50 100 185
24 1.08 25.92 50 1200 1296
19 1.95 37.05 50 950 1852.5
84 1.7 142.8 50 4200 7140
1 6.45 6.45 50 50 322.5
1 10.9 10.9 50 50 545
1 7.21 7.21 50 50 360.5
1 8.01 8.01 50 50 400.5
13.3 2.23 29.659 50 665 1482.95
1 6.48 6.48 50 50 324
154.3 288.33 7715 14416.5
sec 10550N 19865 602068
average assay 30.30797634
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D382 10600 10558 21 2.2 46.2 2.2 50 1050 2310
sum 21 46.2 1050 2310
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RC659 10600 10819 3 1.14 3.42 1.3028 50 150 171
sum
21 1.02 21.42
50 1050 1071
1 7.73 7.73 50 50 386.5
25 32.57 1250 1628.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD391 10600 10832 14 3.93 55.02 4.4514 50 700 2751
sum
35 4.21 147.4
50 1750 7367.5
1 20.2 20.2 0 0
50 222.6 2450 10119
sec 10600N sum 4750 14057
average assay 2.959368421
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D350 10650 10612 2 4.8 9.6 2.2063636 50 100 480
sum
9 1.63 14.67 50 450 733.5
11 24.27 550 1213.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D360 10650 10608 168 2.55 428.4 3.2710545 50 8400 21420
sum
18 3.43 61.74
50 900 3087
10 4.42 44.2 50 500 2210
49 3.95 193.55 50 2450 9677.5
5 10.13 50.65 50 250 2532.5
25 4.84 121 50 1250 6050
275 899.54 13750 44977
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1371 10650 10703 34 1.11 37.74 1.2297143 50 1700 1887
sum
1 5.3 5.3 50 50 265
35 43.04 1750 2152
Hole N E D(m) Au % % * D Avg hole(%) Eff.(w) w*D w*D*%
RCD1279 10650 10818 7 1.25 8.75 9.6926667 50 350 437.5
sum
3 4.6 13.8
50 150 690
3 1.1 3.3 50 150 165
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2 59.77 119.54 50 100 5977
15 145.39 750 7269.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD1279 10650 10618 120 1.63 195.6 1.63 50 6000 9780
sum 120 195.6 6000 9780
sec 10650N sum 22800 65392
average assay 2.868070175
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RC677 10700 10415 1 24.3 24.3 24.3 50 50 1215
sum 1 24.3 50 1215
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD402 10700 10858 8 9.09 72.72 10.733333 50 400 3636
sum
4 14.02 56.08 50 200 2804
12 128.8 600 6440
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD738 10700 10420 16 1.09 17.44 1.09 50 800 872
sum 16 17.44 800 872
sec 10700N sum 1450 8527
average assay 5.880689655
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D351 10750 10590 10 2.18 21.8 3.8235 50 500 1090
Sum
15 3.02 45.3
50 750 2265
2 6.7 13.4 50 100 670
36 2.83 101.9 50 1800 5094
3 6.43 19.29 50 150 964.5
24 1.69 40.56 50 1200 2028
9 8.28 74.52 50 450 3726
1 65.6 65.6 50 50 3280
100 382.4 5000 19118
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D356 10750 10589 18 1.06 19.08 1.433415 50 900 954
Sum
17 1.53 26.01
50 850 1300.5
6 2.28 13.68 50 300 684
41 58.77 2050 2938.5
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D395 10750 10807 1 573 573 573 50 50 28650
Sum 1 573 50 28650
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D648 10750 10490 2 1.6 3.2 1.6 50 100 160
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Sum 2 3.2 100 160
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D757 10750 10485 5 8.34 41.7 3.429032 50 250 2085
Sum
1 33 33
50 50 1650
15 0.9 13.5 50 750 675
10 1.81 18.1 50 500 905
31 106.3 1550 5315
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Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
D1331 10750 10925 1 3.41 3.41 6.03 50 50 170.5
Sum
1 8.65 8.65 50 50 432.5
2 12.06 100 603
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RC678 10750 10464 8 1.16 9.28 3.675 50 400 464
Sum
4 7.45 29.8
50 200 1490
2 14.15 28.3 50 100 1415
6 1.02 6.12 50 300 306
20 73.5 1000 3675
Hole N E D(m) Au (g/t) AU * D Avg hole AU. DIST.BET holes w*D w*D*AU
RCD834 10750 10646 31 2 62 3.045741 50 1550 3100
Sum
4 6.92 27.68
50 200 1384
4 5.77 23.08 50 200 1154
1 20.4 20.4 50 50 1020
13 2 26 50 650 1300
1 5.31 5.31 50 50 265.5
54 164.5 2700 8223.5
sec 10750N 12550 68683
average assay 5.472709163
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calculation of average assay and T.F from sections.
Sec SEC.area(A)m² SEC.INTERVAL av. depth Volume w= V / TF Au(g/t) w * Au
10000 5320 152 446.1 2373252 6787501 4.756606 32285467
10050 5750 115 575.725 3310419 9467798 3.974663 37631301
10100 2700 54 588.65 1589355 4545555 3.000312 13638083
10150 8450 169 532.233 4497371 12862482 2.146457 27608761
10200 7450 149 534.66 3983217 11392001 22.57196 2.57E+08
10250 13550 271 466.45 6320398 18076337 4.866921 87976107
10300 19300 386 444.15 8572095 24516192 5.888369 1.44E+08
10350 5656 113 633.125 3580955 10241531 5.504621 56375750
10400 7450 149 525.9 3917955 11205351 3.724711 41736697
10450 15000 300 566.244 8493660 24291868 3.635055 88302273
10500 18750 375 506.18 9490875 27143903 3.336878 90575886
10550 7150 143 637.175 4555801 13029592 30.30798 3.95E+08
10600 13700 274 315.45 4321665 12359962 2.959368 36577681
10650 10000 200 521.733 5217330 14921564 2.82807 42199230
10700 22150 443 336 7442400 21285264 2.88069 61316240
10750 23050 461 331.104 7631947 21827369 5.472709 1.19E+08
Mine 2.44E+08 6.740961 1.53E+09
Tonnage Factor (T.F) = 1/density = 1/2.41= 0.4149
Total Tonnage = 243954267.9 tons
average grade = 1532079086 / 243954267.9 = 6.28019 (gm/ton) say 6.3 (gm/ton)
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- 99 - | P a g e
Opening up
Introduction
Open pit mining refers to a method of extracting rocks and minerals from the earth by
their removal from pit .
Open pit mines is used when deposit of commercially useful mineral or rock are found
near the surface; that is, where the overburden (surface material covering the valuable
deposit) is relatively thin or the material of interest is structurally unsuitable for
tunneling (as would be the case for sand, cinder, and gravel). For minerals that occur
deep below the surface—where the overburden is thick or the mineral occurs as veins
in hard rock—underground mining methods extract the valued material.
Open-pit mines that produce building materials and dimension stone are commonly
referred to as quarries. People are unlikely to make a distinction between an open-pit
mine and other types of open-cast mines such as quarries, borrows, placers, and strip
mines.
Open-pit mines are typically enlarged until either the mineral resource is exhausted, or
an increasing ratio of overburden to ore makes further mining uneconomic. When this
occurs, the exhausted mines are sometimes converted to landfills for disposal of solid
wastes. However, some form of water control is usually required to keep the mine pit
from becoming a lake.
Generally the mining method in regular ores determines by the stripping ratio that
refers to the ratio between volume of over burden and volume of ore, but in disseminate
cute of grade determine the difference between ore and waste,
Cutoff grade = mining cost / (value X recovery)
The mining method used in this mine is surface mining "selective open pit mining
method because of gold deposit".
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Pit geometry design
Final pit limits that refer to the final limits that ore can be extracted and processed with
a profit affected by:
1. Technical factors,
2. Economic factors.
Pit geometry:
1. Pit overall slope angle,
2. Overall pit depth,
3. Pit width and length.
Pit overall slope angle determination
Slope stability analysis forms an integral part of the opencast mining operations
during the life cycle of the project. In Indian mining conditions, slope design
guidelines were not yet formulated for different types of mining practices and there is a
growing need to develop such guidelines for maintaining safety and productivity. Till
date, most of the design methods are purely based on field experience, rules of thumb
followed by sound engineering judgment. During the last four decades, the concepts of
slope stability analysis have emerged within the domain of rock engineering to address
the problems of design and stability of excavated slopes.
The actual slope angles used in
the mine depend upon:
1. The presence of haulage roads,
or ramps, necessary for the
transportation of the blasted
ore from the pit,
2. Possible blast damage,
3. Ore grades,
4. Economic constraints.
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In slope stability analysis, engineering geological investigation of geological data
collection is necessary to determine the regional geology of the project area and to
obtain the reliable data for proposed project area. There are two investigation stages:
1. Initial site investigation,
2. Final site investigation.
In slope failure analysis, determination of mode of failure is a very important and it
can be considered by the value of slope mass rating (SMR).
Plane failure occurs when a geological discontinuity, such as a bedding plane,
strike parallel to the slope face and dip into the excavation (daylight) at an angle greater
than the angle of friction as shown in Fig6(a).
Wedge failure When two discontinuities strike obliquely across the slope face
and their line of intersection daylight in the slope face, the wedge of the rock resting on
these continuities will slide down the line of intersection, provided that the inclination
of this line significantly greater than the angle of friction as shown in Fig6 (b).
Circular failure When the material is weak (soil) or very heavily fractured (waste
rock dump), the failure will be defined by a single discontinuity surface but will tend to
follow a circular failure as shown in Fig6 (c).
Toppling failure can be occurred when the pit slope angle is greater than the angle
of internal friction and the ratio of block width height is less than the friction angle as
shown in Fig6 (d)
- 112 - | P a g e
Fig 6: Types of failure (a) Plane failure, (b) Wedge failure, (c) Circular failure and (d) Toppling failure.
To prevent all modes of failure and to
achieve safety work the overall slope
angle takes to be 55º.
The inter ramp slope angle takes to be
70 º.
Pit depth
Shallow ore deposits are mined by surface methods but a depth is reached in the case of
most deposits after which underground methods are applied for the extraction of the
remaining ore.
- 113 - | P a g e
From sections:
Section Pit depth
10000mN 221.05
10050mN 225.7
10100mN 318.18
10150mN 323.8
10200mN 333.3
10250mN 360
10300mN 375
10350mN 375
10400mN 350
10450mN 340
10500mN 310
10550mN 324
10600mN 304
10650mN 307
Average pit depth = 295m
Pit width:
Av.Wb = Av. Wd + (ht/tan φ)
Wd: horizontal thickness of the ore body (m),
φ : pit side slope angle along the foot-wall (deg),
ht : pit depth (m).
Wb = 187.319 +2 (295/tan55º)
=450 m
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Pit length:
Lb = Ld + (ht/tan φ)
Ld : total distance between sections (m),
φ : pit side slope angle along the foot-wall (deg),
ht : pit depth (m).
Lb = 650 + (295/tan55º)
=850 m
Parameter value
Over all slope angle 55 º
Pit depth 295 m
Pit width 320 m
Pit length 850 m
- 115 - | P a g e
Bench design
A bench may be defined as a ledge that forms a single level of operation above
which mineral or waste materials are mined back to a bench face. The mineral or waste
is removed in successive layers, each of which is a bench. Several benches may be in
operation simultaneously in different parts of, and at different elevations in the open pit
mine.
Bench elements:
1. Toe,
2. Crest,
3. Floor,
4. High wall (face),
5. Slope angle (angle of inclination) :α,
6. Bench height,
7. Working area,
8. Safety berm.
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Bench parameters
1. Bench height:
The bench height is the vertical distance between each horizontal level of the pit.
The height will depend on :
1. The physical characteristics of the deposit,
2. The degree of selectivity required in separating the ore
3. The size and type of equipment to meet the production requirements,
4. Climatic conditions.
The bench height should be set as high as possible within the limits of the size and
type of equipment selected for the desired production.
The bench should not be so high that it will present safety problems of towering
banks of blasted or un-blasted material.
The bench height in open pit mines will normally range from 15m in large copper
mines to as little as 1 m in uranium mines.
So bench height = 9 m, 3m for every left
2. Bench slope angle:
It's the angle between High wall (face) and the horizontal plane and for granite "hard
rock it's hardness about 7" bench slope angle varies from 70 º to 90 º
α = 5
Because of:
1. Relative small bench height,
2. Hard rock.
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3. Working area:
The width of working bench areas varies within limits depending upon the
equipment used and the mode of operation.
— Wb = 1.2C + Ws + Ws.b + L
Wb : working bench width
C : cut width
Ws : trucks movement "10-15"
L : marginal area "10-20m―
Ws.b = 2 h » h= safty berm height
h= The height of the berm should be of the order of the tire rolling radius,
h=1.5m
Wb = 1.2*18 + 5 + 3 + 5
Wb = 35 m
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4. Bench number:
Bench number = pit averall height / bench height
=295/9
=33 benches
Bench parameters
Bench parameter value
Bench height 9 m
Bench width 35 m
Bench slope angle 5
Number of benches 33
- 119 - | P a g e
FIG 7: Cut sequence for sequential pushbacks
Layout of excavating bench
- 111 - | P a g e
The time sequence showing shovel loading with single spotting.
- 111 - | P a g e
Main haul road design
Over the past three decades, off-highway trucks have been developed for surface
mines up to 400 tons (360 t) in capacity, although 200 tons (180 t) is generally the
largest size in fleet operation. They represent a huge capital investment and a
significant percentage of total costs (Table 13.4.1). If haul road design is inadequate,
the trucks can be highly lethal in the confines of a surface mine. Despite these facts,
haul road design until recently has received little attention.
Haul road design factors must ensure:
1. Minimum costs on a net present value basis for the transport of mineral and
waste throughout the life of the mine.
2. A minimum of traffic congestion and the maintenance of safe, ready access to
the mining operations.
3. The avoidance of areas where slope stability problems could occur.
4. The uses of long-life haul roads rather than short-life roads. This reduces haul
road overall construction costs and operating costs as well as reducing the
demand for haul road construction materials which may not be available in
sufficient quantities from the overburden.
Other factors include the locations of mineral preparation plants, stock yards,
external waste dumps, environmental constraints, etc. All these factors direct attention
to:
1. Haul road layout.
2. Haul road geometry.
3. Haul road construction materials.
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Haul road layout:
For opening up through deposit may layouts of haul road available, like:
1. Switch back road:
They are used to gain altitude or depth without sharp turns.
Fig 8 switch back road layout.
2. Loop road way:
This method is ordinarily employed in mining of deep seated deposits with the use
of motor transport in inclined ingoing trench is driven from the ground surface along
the non-productive side of
open pit at a maximum
permissible gradient at the
end of the open pit field the
road line is turned by 180º.
Fig 8 loop route layout
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3. Spiral route:
The basic conditions for the use of this
method are a level or slightly inclined altitude
of the deposit circular or aver out lines and
considerable size of deposit.
For uphill and downhill pit spiral route
layout used for reaching ore because of relative
small pit geometry.
Haul road geometry:
Number of Lanes:
In-pit roads are usually constructed for single-lane, unidirectional traffic or
two-lane, directional traffic
1. Because traffic density may not be high or
2. Because of space problems. Haul roads from the pit to external waste dumps,
preparation plants, etc.
Number of lanes = 2
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Haul road width:
WR = n (wvehicle+1) +1
Wvehicle: Vehicle width
n: Number of lanes
WR = 2 (4 + 1) +1
Haul road width=11 m
Gradients:
Maximum gradients may be statutorily limited to between 8 to 15% (5 to 8.5°) for
sustained gradients, but in general when considering the economics of uphill haulage,
as well as downhill safety, the optimum gradient for most situations is about 8% (4.5°)
but up to 12% (6.8°) for trolley-assist trucks. For safety and drainage reasons, long
steep gradients should include 150ft (50-m) long sections with a maximum gradient of
2% (1°) for every 1500 to 1800 ft. (500 to 600 m) of severe gradient.
Gradient = 8% (4.5°) to save power and truck maintenance
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Main haul road length
Length of main haul road from pit to waste dump and to processing plant = 2 km.
Reference:
1.SME mining engineering handbook
2. introductory to mining engineering
3. Prof.DR\ Mohamed Abd EL Tawab el gendy
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Drilling and balsting
PART I
Drilling principles
1.1 Introduction:
In virtually all forms of mining, rock is broken through drilling and
blasting. Except in dimension stone quarrying, drilling and blasting are
required in most surface mining. Only the weakest rock, if loosely
consolidated or weathered, can be broken without explosives, using
mechanical excavators (ripper, wheel excavators, shovels etc.) or
occasionally a more novel device, such as a hydraulic jet. In the mining
cycle, drilling performed for the placement of explosives is termed
production drilling. Drilling is also used in surface mining for purposes
other than providing blast-holes. There are minor applications of rock
penetration in surface mining other than drilling. In quarrying,
dimension stone is freed by cutting, channeling, or sawing.
1.2 Classification of methods:
1. Mechanical attack
2. Thermal attack
3. Fluid attack
4. Sonic attack
5. Chemical attack
6. Other methods of attack (electrical, light, or nuclear)
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1.3 PENETRATION RATE:
A function of:
The rock.
The drilling method.
The size & type of bit
The rock properties which effect penetration rate are:
Hardness.
Texture.
Breaking characteristic.
Formation .
1.4 DRILLING METHOD
1.4.1.ROTARY
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1.4.2. ROTARY-PERCUSSION
1.4.3. DOWN HOLE (DH):
The DH drill provides striking energy directly to the bit. There is
rotation so the bit strikes fresh rock with each blow.
Down Hole Drill maintains constant penetration rate at all depths.
Compressed air conducted through the drill steel is used to flush the drill
cuttings from the hole. Performance will not decrease as depth increases.
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PART 2
EXPLOSIVES ENGINEERING
2.1 INTRODUCTION :
The use of explosives in mining and construction applications dates back
to 1627. From 1627 through 1865, the explosive used was black powder.
Black powder was a different type of explosive than the explosives used
today. In 1865, Nobel invented nitroglycerin dynamite in Sweden. He
invented gelatin dynamites in 1866. These new products were more
energetic than black powder and performed differently since
confinement of the explosive was not necessary to produce good results,
as was the case with black powder. From 1867 through the mid-1950's,
dynamite was the workhorse of the explosive industry.
In the mid-1950's' a new product appeared which was called ANFO,
ammonium nitrate and fuel oil. This explosive was more economical to
use than dynamite. During the decades of the 1970's and the 1980's,
ANFO has become the workhorse of the industry and approximately
80% of all explosives used in the United States was ammonium nitrate
and fuel oil.
Other new explosive products appeared on the scene in the 1960's and
1970's. Explosives, which were called slurries or water gels, have
replaced dynamite in many applications. In the late 1970's' a
modification of the water gels called emulsions appeared on the scene.
The emulsions were simple to manufacture and could be used in similar
applications as dynamites and water gels. Commercial explosives fall
into three major generic categories, dynamites, blasting agents and
slurries (commonly called water gels or emulsions).
Blasting problems generally result from poor blast design. Poor
execution in drilling and loading the proposed design and because the
rock mass was improperly evaluated.
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Blast design parameters such as burden, stemming, subdrilling, spacing
and initiation timing must be carefully determined in order to have a
blast function efficiently, safely and within reasonable vibration and air
blast levels.
2.2 SOURCES OF EXPLOSIVE'S ENERGY :
Two basic forms of energy are released when high explosives react. The
first type of energy will be called shock energy. The second type will be
called gas energy.
Figure 2.1 Pressure Profiles for Low and High Explosives
2.2.1 SHOCK ENERGY:
In high explosives, a shock pressure spike at the reaction front travels
through the explosive before the gas energy is released.
The shock energy is commonly believed to result from the detonation
pressure of the explosion. The detonation pressure is a function of the
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explosive density times the explosion detonation velocity squared and is
a form of kinetic energy.
Where:
P = Detonation pressure (Kbar, 1 Kilobar a 14,504 psi)
= Specific gravity of the explosive
= Detonation velocity (fth)
2.2.2 GAS ENERGY:
The gas energy released during the detonation process causes the
majority of rock breakage in rock blasting with charges confined in
boreholes. The gas pressure. often called explosion pressure. is the
pressure that is exerted on the borehole walls by the expanding gases
after the chemical reaction has been completed.
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PART 3
MECHANICS OF ROCK BREAKAGE
3.1 SHOCK ENERGY IN ROCK BREAKAGE :
Unconfined charges placed on boulders and subsequently detonated
produce shock energy which will be transmitted into the boulder at the
point of contact between the charge and the boulder.
3.2 CONFINED CHARGES IN BOREHOLES:
Three basic mechanisms contribute to rock breached with charges
confined in boreholes. The first and least significant mechanism of
breakage is caused by the shock wave.
At most, the shock wave causes microfractures to form on the borehole
walls and initiates microfractures at discontinuities in the burden. This
transient pressure pulse quickly diminishes with distance from the
borehole and since the propagation velocity of the pulse is
approximately 2.5 to 5 times the maximum crack propagation velocity,
the pulse quickly outruns the fracture propagation.
The two major mechanisms of rock breakage results from the sustained
gas pressure in the borehole.
Failure by this mechanism has been recognized for many years and is
commonly called radial cracking (Figure 3.1) .
Figure 2.2 Radial Cracking in Plexiglass
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Direction and extent of the radial crack system can be controlled by the
selection of the proper distance from the borehole to the face (burden)
(Figure 3.2).
Figure 3.2 Influence of Distance to Face on Radial Crack System
The second major breakage mechanism occurs after the radial cracking
has been completed.
3.3 BENCH STIFFNESS :
In most blasting operations, the first visible movement occurs when the
face bows outward near the center. (Figure 2.3).
(Figure 3.4) Axisymmetric Bending Diagram
3.4 EFFECTS OF BLASTHOLE LENGTH :
The rock breakage process occurs in four distinctive steps. As the
explosives detonates. A stress wave moves through the rock uniformly
in all directions around the charge. Radial cracks then propagate
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predominantly toward the free face. After the radial cracking process is
finished. High pressure gases penetrate into the cracks approximately
2/3 of the distance from the hole to the face throughout the radial crack
system. Only after the gas has time to penetrate into the crack systems
are the stresses on the face of sufficient magnitude. To cause the face to
move outward.
The first was to determine the effect of the bench height on the bending
and flexural failure, and the second was to determine the effect of
changing geologic conditions on the movement of the burden itself
(Figure 3.5).
Figure 3.5 Finite Element Model Configurations
3.5 BLASTING PARAMETERS :
In order to compare the model's behavior with that of actual field results,
parameters were chosen so that actual burden movement could be
predicted. The model consisted of a single hole, four inches in diameter
(Figure 3.6).
Figure 3.6 Free Body Diagram for Simulated Condition of Bench
Blasting
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For discussion purposes, four different L/B ratios, 1.2, 2.4, 3.6, and 4.0
will be cosidred .
Figure 3.7 XZ-View of the Deformed Geometry Configuration as LIB
Ratio Changes from 1.2 to 4.0
A close correlation between the finite element model and the rectangular
cross section was observed as hown in Figure 3.9.
Figure 3.9
3.6 GEOLOGICAL EFFECTS ON DISPLACEMENT :
In order to analyze the significance of beds of different materials on
bench blasting, In order to analyze the significance of beds of different
materials on bench blasting, five different models were analyzed using
the same finite element model (Figure 3.11).
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Figure 3.11 Geologic Structure of Models
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PART 4
EXPLOSIVE PRODUCTS
4.1 ENVIRONMENTAL CHARACTERISTICS OF EXPLOSIVES :
The selection of the type of explosive to be used for a particular task is
based on two primary criteria. The explosive must be able to function
safely and reliably under the environmental conditions of the proposed
use, and the explosive must be the most economical to use to produce the
desired end result.
4.1.1 SENSITIVENESS :.
4.1.2 WATER RESISTANCE :
4.1.3 FUMES :
4.1.4 FLAMMABILITY :
4.1.5 TEMPERATURE RESISTANCE :
4.1.6 COLD RESISTANCE :
4.1.7 VELOCITY :
4.1.8 DETONATION PRESSURE :
4.1.9 DENSITY :
Where:
De = Diameter of explosive (in)
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4.1.10 STRENGTH :
4.1.11 COHESIVENESS :
4.2 COMMERCIAL EXPLOSWES :
The products used as the main borehole charge can be broken into three
generic categories. Dynamite, slurries and blasting agents.
4.2.1 DYNAMITE :
Dynamites are the most sensitive of all the generic classes of explosives
used today.
4.2.2 GELATIN DYNAMITE :
Gelatin dynamite, used in commercial applications, can be broken into
three subclasses, straight gelatin, ammonia gelatin and semigelatin
dynamites.
4.2.3 SLURRY EXPLOSIVES
Slurry explosive is a mixture of ammonium nitrate or other nitrates and
fuel sensitizers which can either be a hydrocarbon or hydrocarbons and
aluminum. In some cases explosive sensitizers, such as TNT or
nitrocellulose, along with varying amounts of water are used (Figure
4.5). An emulsion is somewhat different from a water gel or slurry in
characteristics, but the composition contains similar ingredients and
functions similarly in the characteristics, but the composition contains
similar ingredients and functions similarly in the blast hole (Figure 4.6).
In general, emulsions
have a somewhat higher
detonation velocity and
in some cases, may tend
to be wet or adhere to
the blast hole causing
difficulties in bulk
loading. For discussion
purposes, emulsions
and water gels will be treated under the generic family of slurries.
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PART 5
PRIMER AND BOOSTER
5.1 PRIMER AND BOOSTER SELECTION :
The difference between a primer and booster is in its use. Rather than in
its physical composition or makeup. A primer is defined as an explosive
unit which contains an initiator, a booster is used to put additional energy
into a hard or tough rock layer (Figure 5.1).
Figure 5.1 Primer and Booster in Borehole
5.2 SELECTION CRITERIA FOR PRIMER :
The two most critical criteria in primer selection are primer composition
and primer size.
5.3 BOOSTER :
Boosters are used to intensify the explosive reaction at a particular
location within the explosive column. In general, boosters are used to put
more energy into a hard layer within the rock column.
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PART 6
BLAST DESIGN
The design of blasts must encompass the fundamental concepts of ideal
blast design, which are then modified when necessary to account for
local geologic conditions. In order to evaluate a blasting plan, the plan
must be taken apart and each variable or dimension must be evaluated. A
plan must be designed and checked one step at a time. This chapter will
lay out a step-by-step procedure for the analysis of a blasting plan.
Methods to determine whether design variables are in normally
acceptable ranges will be examined.
6.1 BURDEN :
Burden distance is defined as the shortest distance to relief at the time the
hole detonates (Figure 6.1).
The selection of the proper burden is one of the most important decisions
made in any blast design. Of all the design dimensions in blasting. It is
the most critical.
Figure 6.1 Symbols for Blast Design
Where:
B = Burden
T = Stemming
J = Subdrilling
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L = Bench height
H = Blasthole depth
PC = Powder column length
If the operator has selected a burden and used it successfully for a drill
hole of another size and wants to determine a burden for a drill hole that
is either larger or smaller, one can do so quite easily if the only thing that
he is changing is the size of the hole and the rock type and explosives are
staying the same. To do this, one can use the following simple ratio:
Where:
B1 = Burden successfully used on previous blasts
De1 = Diameter of explosive for B1
B2 = New burden
De2 = New diameter of explosive for B2
6.1.1 ADJUSTMENTS FOR ROCK & EXPLOSIVE TYPE :
When an operator is moving into a new area where he has no previous
experience, he would have only general rock and explosive
characteristics to work with. When moving into a new area, especially
one where there are residents nearby, it is essential that the first shot not
be a disaster. To approximate burden under these situations, the
following empirical formula is helpful.
[
]
Where:
B = Burden (ft)
SGe = Specific gravity of explosive
SGr = Specific gravity of rock
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De = Diameter of explosive (in)
The previous equations proposed for burden selection used the specific
gravity of the explosives as an indicator of energy. The new generation
of slurry explosives called emulsions have somewhat different energies
but near constant specific gravity. The equation that uses relative energy
is:
√
Where:
B = Burden (ft)
De = Diameter of explosive (in)
Stv = Relative bulk strength (ANFO = 100)
SGr = Specific gravity of the rock
6.1.2 GEOLOGIC CORRECTION FACTORS :
No one number will suffice as the exact burden in a particular rock type
because of the variable nature of geologyTo estimate the deviation from
the normal burden formula for unusual rock structure, two constants are
incorporated into the formula. Kd is a correction for the rock deposition
and Ks is a correction for the geologic structure. Kd values range from
1.0 to 1.18 and describe the dipping of the beds (Table 6.3).
TABLE 6.3 CORRECTIONS FOR ROCK DEPOSITION
BEDDING ORIENTATION Kd
Bedding steeply dipping into cut 1.18
Bedding steeply dipping into face 0.95
Other cases of deposition 1.00
The correction for the geologic structure takes into account the fractured
nature of the rock in place, the joint strength and frequency as well as
cementation between layers of rock. The correction factors for rock
structure ranges from 0.95 to 1.30 (Table 6.4). Massive intact rock
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would have a Ks value of 0.95 while heavily broken fractured rock could
have a Ks value of about 1.3.
TABLE 6.4 CORRECTIONS FOR GEOLOGIC STRUCTURE
GEOLOGIC STRUCTURE Ks
Heavily cracked, frequent weak joints weakly cemented
layers
1.30
Thin well-cemented layers with tight joints 1.10
Massive intact rock 0.95
6.2 STEMMING :
The common material used for stemming is drill cuttings, since they are
conveniently located at the collar of the blast hole.
T = 0.7 x B (For crushed stone or drilling chips)
Where:
T = Stemming (ft)
B = Burden (ft)
Figure 6.2 Stemming Zone Performance
6.3 SUBDRILLING :
Subdrilling is a common term to denote the depth which a blasthole will
be drilled below the proposed grade to ensure that breakage will occur to
the grade line. Blastholes normally do not break to full depth.
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J= 0.3×B
Where:
J = Subdrilling (ft)
B = Burden (ft)
Figure 6.5 Problems of Soft Seam off Bottom
5.4 BLASTING CONSIDERATIONS :
The blasting consideration of fragmentation, air blast, flyrock and
ground vibration would have to be assessed.
The more massive the rock in a production blast. The more probable the
outcome listed in Table 5.5.
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TABLE 5.5 POTENTIAL PROBLEMS AS RELATED TO
STIFFNESS RATIO (L/B)
6.5 SPACING :
If holes are initiated simultaneously. Spacing's must be spread further
apart than if holes are timed on a delay. (Figure 6.6).
Figure 6.6 Shattered Zones from Close Spacing
Figure 6.7 Rough Walls from Excessive Spacing
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6.5.1 INSTANTANEOUS INITIATION LOW BENCHES
In order to check the blasting plan and determine if spacing is within
normal limits, the following equation can be used:
Where:
S= Spacing (ft)
L = Bench height (ft)
B = Burden (ft)
6.5.2 INSTANTANEOUS INITIATION HIGH BENCHES
6.5.3 DELAYED INITIATION LOW BENCHES :
When the stiffness ratio is between one and four with delayed initiation
between holes, the following relationship is used to check spacing:
Where:
S = Spacing (ft)
L = Bench height (ft)
B = Burden (ft)
6.5.4 DELAYED INITIATION HIGH BENCHES :
S=1.4 B
Where:
S = Spacing (ft)
B = Burden (ft)
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If the calculated spacing value is within plus or minus 15% of the actual
spacing, the spacing is within reasonable limits.
6.6 Drilling pattern:
Figure 6.12 V-Cut (Square Corner), Progressive Delays, S = 1.4 B
Figure 6.12 Stagger pattern
6.7 Initiation pattern:
There are two types of initiation patterns
6.7.1 Chevron initiation pattern (closed)
6.7.2 Corner initiation
pattern (flat)
Corner initiation pattern is
used because it gives good
control of rock
displacement and gives
good fragmentation
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PART 7
CALCAULATION
7.1 BURDEN :
√
We take
√
=15.15ˋ
=4.55 m
=4.55×1×0.95
=4.5 m
7.2 STEMMING :
=0.7×4.5
=3.15 m
7.3 SUBDRILLING :
=0.3×4.5
=1.35 m
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7.4 STIFFNESS RATIO :
We take L= 9 m
= 2
7.5 SPACING :
Delayed initiation with the S.R greater than 1 but less than 4
= 5 m
7.6 CHARGE LENGTH :
=7.2 m
7.7The total weight of explosive per Colum :
› loading density
=132.3 kg
7.8VOLUME UNDER AREA :
= 202.5 m
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7.9 POWDER FACTOR :
⁄
7.10Weight of rock broken per hole :
= 570 t
The product of ore per day is 10000 t.
The stripping ratio is 1:5.
The product of broken rock per day is 60000 t.
7.11 The number of hole:
= 105 Hole
We use two face.
For one face length is 65 m and width is 18 m.
For one face 4 rows and 13 Colum.
7.12 THE NUMBER OF DRILLS:
Rate of drilling is 35 m/hour.
Total meters drilled per shift /drill =35 x 5 =175 m/shift
No. of drills
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We use 3 drills
We work 2 shifts per day (2 shift/day)
7.13 The COST OF DRILLING AND BLASTING:
The cost of one cubic meter of drilling and blasting is about 6:10 LE
(We take the cost is ( 8 ⁄
Cost
7.14 CHAPTER 7 SUMMARY
type of explosive is emulsion
1 BURDEN (B) M 4.5
2 STEMMING (T) M 3.15
3 SUBDRLLING (J) M 1.35
4 STIFFNESS RATIO (S.R) 2
5 SPACING (S) M 5
6 CHARGE LENGTH ( ) M 7.2
7 CHARGE PER HOLE ( ) kg 132.3
8 POWDER FACTOR (P.F)
⁄
0.65
9 NUMBER OF HOLE (n) 105
10 ORE PRODUCTION T 10000
11 TOTAL CHARGE ( ) kg 13891
12 NUMBER OF DRILLS 3
13 COST (LE) 168480
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References:
Blasters‘ Handbook, E. I. du Pont de Nemours & Co., Pont de
Nemours & Co., Wilmington, DE W
Rock Blasting and Overbreak Overbreak Control, National
Highway Control, National Highway Institute, Washington, DC
Institute, Washington, DC
―Introductory mining engineering ―, Howard l. Hartman &
JanM.mutmansky, Scound Edition ,new York .2002.
Lectures by Prof. Dr. Mohamed Abd El Tawab El Gendy
Lectures by DR. Ibrahim assakkaf
Lectures by eng. Abd-Elmoneam seleim
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Loading and transportation
Introduction
Equipment selection is one of the most important steps of open pit design. Mining
costs are mainly affected by the number and capacity of equipment.
Equipment selection effects economic considerations in open-pit design,
especially overburden or (waste rock) and ore mining costs and cost increasing
parameters as a function of plan location and depth. Mining costs are a function of site
conditions, operating scale and equipment. The purpose of equipment selection is to
select optimum equipment with minimum cost.
Equipment selection for open-pit mines is a very important decision which will
impact greatly the economics operations.
Selection of loading equipment
The three main types of surface loading equipment:
1. Cable shovels,
2. Hydraulic shovels,
3. Front end loaders.
The following table shows comparison between the Operating
characteristics of each type:
Cable Shovel Hydraulic Shovel Front-end Loader Proven and reliable, low
operating costs and machine
availability.
Significant disadvantages in
un-blasted toes.
No selectivity; high blending
vertically.
Cannot dig below floor level.
Cannot traverse gradients
greater than 1 in 20.
No control over dump rate.
Has long life (20-30 years).
Low operating costs
compared to hydraulic
shovels or FELs.
Powerful digging force permits
the machine to operate in a variety
of conditions.
Half the amount of hoist and
crowd needed for same size
bucket as a cable shovel.
Can be used as back hoe and can
dig below the floor.
Half the size of shovel for same
bucket size as cable shovels.
Double speed of cable shovel.
Very high selectivity.
Horizontal crowd motion can load
thin seams.
Discharge into trucks can be
controlled.
Versatile, highly mobile,
multipurpose.
Increasingly used as LHD machines.
Traction and stability problems.
Bucket wider than front tires (tire
protection).
Tire costs considerable.
Not suitable for soft grade. High
maneuvering required, increasing risk
of overturning or falling off edge.
Well suited for: well fragmented
ground on good under-footing, road
maintenance, stockpile reclamation
and flexible back-up machine.
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Characteristics of local operating condition in the proposed mine:
badly fragmented product (R.O.M) on bad under-footing.
Front-end Loader is no longer useful.
High selectivity is needed to load the ore and the waste separately.
Electric Cable Shovel is no longer useful.
Thus,for more selectivity we have to use proper sized Hydraulic Shovel as back
hoe
Selection of haulage equipment
The three main types of surface haulage equipment which are generally
available for mine use are:
1. Belt Conveyor Haulage.
2. Locomotive Haulage.
3. Truck Haulage.
The following table shows comparison between the Operating characteristics
of each type:
Equipment Belt Conveyor Haulage Locomotive Haulage Truck Haulage
Operation(output) Continuous (not cyclic) Cyclic (discontinuous) Cyclic (discontinuous)
Capacity (t/hrs.) High high Moderate
Haulage distance
(km)
Limited (up to 5 km) unlimited Limited (0.5 – 10 km)
Relief Moderate low Moderate
Grade High (Up to 20o) Low (up to 3
o) Moderate (8
o – 12
o)
Advantages High output at relative
low cost.
Better grade ability
High daily output.
Long haulage distance.
Coarse / blocky product.
High flexibility.
High maneuverability.
Moderate grade
ability.
Coarse product.
Disadvantages Low flexibility.
High installation cost.
Limited to fine product.
Poor grade ability.
High installation cost.
High operating cost
(esp. with bad roads)
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Characteristics of local operating condition in the proposed mine:
The grade is moderate.
Locomotive Haulage must be excluded.
The daily Capacity is moderate about (10,000 t/sh.), haulage distance is
moderate (2 km), high flexibility and maneuverability is needed for advanced
face.
Belt Conveyor Haulage must be excluded.
Thus, for more flexibility and maneuverability we have to use Truck Haulage.
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Estimation of proper size and number of Loading and haulage
equipment
Design and operating data:
Ore: gold with a cut of grade equal 0.5 g/t
Associated rock is granite, well blasted
Required production of ore = 10000 t/day
Waste associated with ore, 50000 t/day (source from mine field)
Total required production = 60000 t/day
Two faces (two shovels) are selected to cover the required production
Working time:3 shift/day, 8 hours / shift, effective time of loading and
haulage 6 hours/shift
Haulage distance to crushing plant = 2000 m
Haulage distance to stock pile = 2000 m
Job condition and management is considered to be good
Loading equipment selection
Proposed production per face = 60000/2=30000 t/day (One face)
Proposed production per shift = 30000/3=10000 t/sh. (One face)
Cycle time range of loader = 28- 45 sec,
We consider as an average cycle time=35 sec
SO No. of cycles per hour (C) = 3600/35 = 102cycles/hr. and the actual
bucket size used calculated by:
BC = [Q (m3 bank/ hr) * swell factor] / [C * S * O.A * BF]
Where:
(BC)=bucket capacity (m3)
(Q)= productivity (m3 bank/hr.)
(C)= no. of cycles per hours [100 cycle/hr.]
(S)=swing factor
(O.A)=operating efficiency
(BF)=bucket fill factor
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From table [13.3.3 SME handbook] @ angle of swing = 60 o
Swing factor (s) = 1.1
From table [13.3.4 SME handbook] @ good job and good
management conditions
Operating efficiency (O.A) =0.73
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From table [13.3.5 SME handbook] @ granite material type
Bucket fill factor (fillability) (BF) = 0.8
Bank density (ρbank) =2.41 t/ m3
Swell factor = 1.55
Loose density (ρloose)=2.41/1.55=1.55 t/m3
Shovel production per hour (t/hr)=10000/6=1666.67 t/hr
Shovel production per hour (m3 bank /hr) = (t/hr)/ ρbank=1666.67 /2.41=691.56 m
3 bank/hr
BC= [Q (m3 bank/ hr) * swell factor] / [C * S * O.A * Bf]
BC = (691.65 * 1.55) / (102 * 1.1 * 0.73 * 0.8) = 16.36 m3
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The nearest mining excavator size of bucket (17 m3), we will use one back
hoe (for more selectivity) hydraulic excavator RH 120 E (TEREX) with
bucket size (17m3)
Actual productivity of loader = BC * C * S * O.A * Bf * ρloose =17 * 102*
1.1*0.73*0.8*1.55= 1726.57 t/hr
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Actual productivity of loader (t/sh.) =1726.27*6=10359.47 t/sh.
(Satisfy face shift production)
Daily production of one face = 10359.47*3=31078.4 t/day
Total daily production of the two faces =31078.4*2=62156.8 t/day
Haulage equipment selection;
Bucket capacity per one pass = BC * BF * ρloose
=17*0.8*1.55
=21.08 t/pass
Where:
BC=bucket capacity (m3),
BF=bucket fill factor,
ρloose= the loose density of material.
To fill the truck we use 4:6 passes:
In case of 4 passes:
Truck capacity=21.08*4=84.32 t
Difference (%) = (90.9-84.32)/90.9=7.2%
In case of 5 passes:
Truck capacity=21.08*5= 105.4t
Difference (%) = (109-105.4)/109=3.3%
In case of 6 passes:
Truck capacity=21.08*6=126.48 t
Difference (%) = (136-126.48)/136=7%
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So we use truck capacity of (109) [MT 3000] and fill it with (5) passes
Truck cycle time = loading time + haulage time (loaded) + dumping time +
returning time (empty) +spotting time
From [table 9.3.4 SME handbook] @ average conditions and Rear-dump truck
type
Dumping time =1.3 min
From [table 9.3.5 SME handbook] @ average condition and Rear-dump truck
type
Spotting time = 0.3 min
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Loading time = no. of passes * loader cycle time =5*35=175 sec=3 min
Haulage distance (Loaded) is about 2000 m is divided into :-
200 m face @ speed = 15 km/hr
200 m ramp @ speed = 25 km/hr
1400 m main road @ speed = 40 km/hr
200 m dump area @ speed = 15 km/hr
Returning distance (Empty) is about 2000 m is divided into :-
200 m face @ speed =20 km/hr
200 m ramp @ speed = 30 km/hr
1400 m main road @ speed = 45 km/hr
200 m dump area @ speed = 20 km/hr
Hauling time (loaded) =
200/ (1000*15) +200/ (1000*25) +1400/ (1000*40) +200/ (1000*15) = 0.0697
hr=4.182 min
Returning time (empty) =
200/ (1000*20) +200/ (1000*30) +1400/ (1000*45) +200(1000*20)
=0.058hr=3.47 min
Truck cycle time = 3+4.182+1.3+3.47+0.3=12.25 min
No. of truck cycles per shift = (60*6)/12.25= 29 cycle/shift
Productivity of one truck /sh. =105.4*29=3056.6 t/sh.
Req. number of trucks= (loader productivity /sh.)/ (one truck
productivity/sh.)=10359.47/3056.6= 4 trucks
Actual required no. of truck =4/0.83=5 trucks
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Synchronization of loader and trucks
n ( tspotting + tloading ) >= tcycle
4(0.3+3)>=12.25
13.2>12.25
Difference = 13.2-12.25=0.95 min
Reference
1. SME mining engineering hand book
2. Introductory to mining engineering
3. Prof.Dr\ Mohamed abd el tawab el gendy's lectures
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Ore dressing
2.1.Crushing
Crushing is the first mechanical stage in the process of comminution in which the
main objective is the liberation of the valuable minerals from the gangue.
2.2 .Primary crushers
Primary crushers are heavy-duty machines, used to reduce the run-of-mine ore down
to a size suitable for transport and for feeding the secondary crushers or AG/SAG
mills. They are always operated in open circuit, with or without heavy-duty scalping
screens (grizzlies). There are two main types of primary crusher in metalliferous
operations –jaw and gyratory crushers- although the impact crusher has limited use
as a primary crusher and will be considered separately.
2.3 .Jaw Crusher
Jaw Crusher is one of the main types of primary crushers in a mine or ore processing
plant. The size of a jaw crusher is designated by the rectangular or square opening at
the top of the jaws (feed opening). For instance, a 24 x 36 jaw crusher has a opening
of 24" by 36", a 56 x 56 jaw crusher has a opening of 56" square. Primary jaw
crushers are typically of the square opening design, and secondary jaw crushers are
of the rectangular opening design. However, there are many exceptions to this
general rule. A Jaw Crusher reduces large size rocks or ore by placing the rock into
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compression. A fixed jaw, mounted in a "V" alignment is the stationary breaking
surface, while the movable jaw exerts force on the rock by forcing it against the
stationary plate. The space at the bottom of the "V" aligned jaw plates is the crusher
product size gap, or the size of the crushed product from the jaw crusher. The rock
remains in the jaws until it is small enough to pass through the gap at the bottom of
the jaws
2.4 .Grinding
Grinding is a powdering or pulverizing process using the rock mechanical forces of
impaction, compression, shearing and attrition.
Milling, sometimes also known as fine grinding, pulverising or comminution, is the
process of reducing materials to a powder of fine or very fine size. It is distinct from
crushing or granulation, which involves size reduction to a rock, pebble or grain size.
Milling is used to produce a variety of materials which either have end uses
themselves or are raw materials or additives used in the manufacture of other
products.
A wide range of mills has been developed for particular applications. Some types of
mills can be used to grind a large variety of materials whereas others are used for
certain specific grinding requirements. This brief aims to present the factors to
consider when choosing a particular grinding applications and to give an overview of
the equipment which is available.
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The two main purposes for a grinding process are:
• To liberate individual minerals trapped in rock crystals (ores) and thereby
open up for a subsequent enrichment in the form of separation.
• To produce fines (or filler) from mineral fractions by increasing the specific
surface.
2.5. Ball Mill
Ball mills are similar in concept to the rod mill but are charged with steel balls in
place of the rods. The mill consists of a cylindrical drum, sometimes tapered at one
end, and usually has a charge of steel balls (up to 40% by volume) ranging in size up
to
125mm for larger mills.
Product size can be as
small as 0.005mm, but
product size is dependant
upon the time the charge
spends in the grinding zone
and therefore the reduction
rate is a function of the throughput. The lining material is of great importance as
there is a significant amount of wear taking place due to the action of the steel balls.
The speed of rotation is optimum at about 75% of critical speed. Some mills are
compartmentalized with each subsequent section having a smaller ball size. The
mineral can pass through to the proceeding section, but the balls cannot. This ensures
that the smaller particles are attacked by the smaller grinding media. It is a versatile
grinding mill and has a wide range of applications. The mill can vary in size from
small batch mills up to mills with outputs of hundreds of tonnes per hour. They are
the most widely used of all mills. Small hand operated ball mills are used in Bolivia
for preparation of ore, sand and
gravel.
3.1. Ore characteristics
Ore competency and hardness is
the prime determinant of the circuit
configuration. Low competency
may permit the use of lower capital
cost crushing and milling
equipment such as MMD sizers,
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single-stage primary mills and SAG mills. High competency will dictate the
examination of SABC circuits and, for ores that exhibit extreme resistance to
breakage, staged crushing (two or three stages) followed by either ball or SAG
milling may have to be considered. At higher throughputs, such circuits can consider
the use of HPGRs in place of SAG milling. Circuits handling moderately competent
ores can take advantage of operating cost savings by employing fully autogenous
grinding (AG) mills.
3.1.Free-Milling Ore Process Options
The recovery circuits of choice are either carbon-in-leach (CIL) or heap leach
followed by carbon-in-solution (CIS) . For free milling ores exhibiting a high gold
recovery at a reasonably coarse grind size and with average oxygen and cyanide
consumptions, the engineer is faced with few selection issues in determining the
flowsheet.
3.2. Gravity-recoverable gold
The benefits of gravity recovery can be readily assessed by undertaking leach tests
with and without pre-treatment coupled with mineralogical analysis and examination
of gold leach tails solids. Early equipment included shaking tables, spirals, drums
and other devices. The alternatives expanded to more sophisticated and efficient
centrifugal separators such as Knelson and Falcon concentrators with the latter
perhaps being more applicable to the treatment of finer solids. In recent times, a
further range of gravity equipment has been successfully commercialized including
in-line pressure jigs. equipment development has also extended to the use of
high-intensity cyanidation devices to solubilize gold from the concentrates.
3.3. Complex Ore Process Options
Complex ores are intermediate between free-milling and refractory ores. As such,
they give rise to high usages of cyanide and oxygen and/or are pregrobbing.
Speciation of cyanide complexes within leach liquors especially for feeds known to
contain copper, zinc, thiosulfates and other complexes is recommended. Flowsheet
selection issues are discussed under the following headings.
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4.1. Methods of treatment gold ore:
1. Gravity concentration
2. leaching
3. Flotation
4.1.1.Gravity concentration
Gravity separation, one of the oldest separation techniques, has become increasingly
popular in modern plants, with new equipment enhancing the range of separations
possible (Laplante and Doucet, 1996). When coupled with generally low capital and
operating costs and lack of chemicals to cause environmental concerns, this often
provides an attractive process for the recovery of gold. Gravity separation relies
upon the differences in density of minerals to provide efficient separation. The ease
and efficiency of separation is dependent on a number of factors, including relative
density, particle size and shape, liberation – all of which affect the selection of
equipment type.
In the case of gold, gravity tools can be useful in solving a number of problems.
These can include what is termed spotty or coarse gold, which makes mass balancing
and gold accounting extremely difficult. By utilizing gravity ahead of the leach train,
early recovery of gold in the process can also have financial benefits and avoid
potential losses. Gravity recovery is also a useful diagnostic tool and can, and has
been, used to check for the potential salting of samples. Removal of coarse
gravity-recoverable gold can also enhance leach kinetics in plant practice. Use of
gravity recovery as a safety net on tailings has also been exploited at several
operations, where unleached gold, either ascoarse particles or sulfide locked, are
recovered from the tailings by gravity
means and re-treated, usually with a re-grind prior to a re-leach. The range of
equipment available for gravity separation includes standard mineral jigs, Kelsey
jigs, In-Line Pressure jigs, spirals, tables, Mozeley sizer, Knelson, Falcon Superbowl
and others.
1.1. Conventional jigs
Conventional jigs are often used to recover heavy minerals that are liberated at a
coarse particle size from crushing/grinding circuits, thus avoiding subsequent
over-grinding and loss.
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1.2. Centrifugal jigs
Centrifugal jigs use enhanced forces generated by their spinning motion to enable
finer particle sizes and closer specific gravity (SG) minerals to be separated. The
Kelsey jig is the most common example of this type of separator.
1.3. Spirals
Spirals are one of the oldest gravity separators. There is a wide range of profiles
available including low-grade, medium-grade, high-grade and fine mineral models,
plus ones incorporating different wash water techniques.
Careful monitoring and control of size distribution is important in achieving
optimum results with spirals.
1.4. Mozley gravity separator (MGS)
The MGS is a low-capacity high-performance gravity separator suitable for treating
difficult fine particle feeds below 75 mm.
1.5. Falcon and Knelson concentrators
These are centrifugal type gravity separators also suited to fine particle-size feeds
(Ancia et al., 1997). These units come in batch and continuous configuration for both
laboratory testing and operational application.
1.6. Shaking tables
Tables are often used in the laboratory as a preliminary test to ascertain an ore‘s
amenability to gravity separation or upgrade, or as a tool in their own right. Size of
tables used in the laboratory environment vary but usually range from third or
quarter production size up to half and on occasion full size.
1.7. Super-panners
A laboratory diagnostic tool used to produce the highest concentration of material,
super-panners are often used in conjunction with other gravity devices as a final
cleaning step, and in the preparation of samples for mineralogical work.
4.2.Leaching
Oxides are leached with a sulfuric acid or sodium carbonate solvent, while sulfates
can be leached with water or sulfuric acid. Ammonium hydroxide is used for native
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ores, carbonates, and sulfides, and sodium hydroxide is used for oxides. Cyanide
solutions are a solvent for the precious metals, while a sodium chloride solution
dissolves some chlorides. In all cases the leach solvent should be cheap and
available, strong, and preferably selective for the values present.
Leaching is carried out by two main methods: simple leaching at ambient
temperature and atmospheric pressure; and pressure leaching, in which pressure and
temperature are increased in order to accelerate the operation. The method chosen
depends on the grade of the feed material, with richer feed accommodating a costlier,
more extensive treatment.
4.2.1.Type of leaching :
1-Leaching in-place, or in situ leaching, is practiced on ores that are too far
underground and of too low a grade for surface treatment. A leach solution is
circulated down through a fractured ore body to dissolve the values and is then
pumped to the surface, where the values are precipitated.
2- Heap leaching is done on ores of semilow grade--that is, high enough to be
brought to the surface for treatment. This method is increasing in popularity as larger
tonnages of semilow-grade ore are mined. The ore is piled in heaps on pads and
sprayed with leach solution, which trickles down through the heaps while dissolving
the values. The pregnant solution is drained away and taken to precipitation tanks.
3-Higher-grade ores are treated by tank leaching, which is carried out in two ways.
One method is of very large scale, with several thousand tons of ore treated at a time
in large concrete tanks with a circulating solution. In the second method, small
amounts of finely ground high-grade ore are agitated in tanks by air or by mechanical
impellers. Both solutions pass to precipitation after leaching is completed.
Pressure leaching shortens the treatment time by improving the solubility of solids
that dissolve only very slowly at atmospheric pressure. For this process autoclaves
are used, in both vertical and horizontal styles. After leaching, the pregnant solution
is separated from the insoluble residue and sent to precipitation.
4.2.2.What is heap leaching?
To those of us in the gold industry, the question ‗‗What is Heap Leaching?‘‘ seems to
have an obvious answer. In the simplistic sense, heap leaching involves stacking of
metal-bearing ore into a heap on an impermeable pad, irrigating the ore for an
extended period of time (weeks, months or years) with a chemical solution to
dissolve the sought-after metals, and collecting the leachant (pregnant solution) as it
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percolates out from the base of the heap. Fig. 2 is an aerial photograph showing the
typical elements of a preciousmetals heap-leach operation – an open-pit mine, a heap
of crushed ore stacked on a plastic pad, ponds, a solution process facility for
recovering gold and silver from the pregnant solution, and an office facility. For a
small operation such as the one illustrated here, very limited infrastructure is
required. In a more complex sense, heap leaching should be considered as a form of
milling. It requires a non-trivial expenditure of capital, and a selection of operating
methods that trade off cost versus marginal recovery. Success is measured by the
degree to which target levels and rates of recovery are achieved. This distinguishes
heap leaching from dump leaching. In dump leaching, ores are stacked and leached
in the most economical way possible, and success is achieved with any level of net
positive cash flow. The bibliography of precious metals heap-leaching is quite
extensive, but a limited bibliography has been compiled.
4.2.3. Why select heap leaching as the processing method?
Gold and silver can be recovered from their ores by a variety of methods, including
gravity concentration, flotation and agitated tank leaching. Methods similar to heap
leaching can be employed: dump leaching and vat leaching (vat leaching is the
treatment of sand or crushed ore in bedded vats with rapid solution percolation).
Typically, heap leaching is chosen for basic financial reasons – for a given situation,
it represents the best, or safest, return on investment. Some of the financial
considerations that might result in the selection of heap leaching are presented
below.
4.2.7. Heap leaching of gold and silver ores
Heap leaching of gold and silver ores is conducted at approximately 120 mines
worldwide. Heap leaching is one of several alternative process methods for treating
precious-metal ores, and is selected primarily to take advantage of its low capital cost
relative to other methods. Based on surveys of about 60% of the known heap-leach
operations, it is likely that heap leaching produces 12% of the world‘s gold. In 2004,
at least 10 major heap leaches were in the late design stages, in such diverse locations
as Brazil, Kazakhstan, Laos, Mexico, Peru, the United States and Uzbekistan. Heap
leaching for silver is conducted using the same principles and operating practices as
for gold, but heap-leach operations produce only a small fraction of world silver
production.
Heap leaching had already become a fairly sophisticated practice at least 500 years
ago. illustrates a heap leach with a 40-day leach cycle which could pass in many
ways for a modern heap leach. The Agricola heap-leach recovered aluminum
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(actually alum) for use in the cloth-dying industry. Copper heap and dump leaches in
southern Spain were common by about 1700. Gold and silver heap-leaching began
with the first Cortez heap leach in 1969. While many projects have come and gone,
Cortez is still going – their new 63,000 t/d South Area leach started up in 2002.
The largest U.S. precious-metal heap-leach is the Round Mountain operation in
Nevada, with over 150,000 t/d of ore going to crushed or run-of-mine heaps, at an
average grade of 0.55 g/t (this chapter follows the convention of ton for short ton and
tonne for metric ton; t/d reflects metric tonnes). Worldwide, Newmont‘s Yanacocha
operation in Peru holds the record, with a production rate of 370,000 t/d, at an
average total reserve gold grade of 0.87 g/t. On the other end of the scale, some very
high-grade ores – up to 15 g/t (0.5 oz/ton) – are being successfully processed at rates
of several hundred tonnes per day (Sterling, Nevada; Hassai, Sudan; Ity, Ivory
Coast). Nevada was the ‗birthplace‘ of modern gold heap-leaching in the late 1960s,
and is only now giving up its dominance of this technology. Other very large gold
districts – notably the pre-Cambrian shield areas of Canada, Australia and South
Africa – show relatively few heap leaches. There are several reasons for this
geographic concentration, but the primary reason is that Nevada gold deposits tend
to have been created by low-energy geologic
4.2.8.Chemistry of gold and silver heap leaching
The chemistry of leaching gold and silver from their ores is essentially the same for
both metals, and many ores contain a mixture of the two. A dilute alkaline solution of
sodium cyanide dissolves these metals without dissolving many other ore
components (copper, zinc, mercury and iron are the most common soluble
impurities). Solution is maintained at an alkaline pH of 9.5 to 11. Below a pH value
of 9.5, cyanide consumption is typically high. Above a pH value of 11, metal
recovery decreases.
Silver is usually not as reactive as gold with cyanide. This is because gold almost
always occurs as the metal, whereas silver may be present in the ore in many
different chemical forms, some of which are not cyanide-soluble. Gold recovery
efficiency from operating heap leaches is typically _70%, although it can range from
50 to 90%. Silver recovery efficiency is typically _55%. Other leaching agents, such
as thiosulfate, thiourea, hypochlorite and bromine, have been experimented with an
alternative to cyanide, but cyanide is by far the most effective and the most
environmentally friendly leaching agent.
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4.2.9.Cyanide in Gold Extraction
As part of an overall planning procedure, the role cyanide will play in a mining
operation must be well understood and defined. Personnel who have a sound
knowledge of the types of cyanide and how their chemistry relates to extraction,
recycling, workers' health and environmental risks should participate in planning the
mine and its cyanide strategy.
1. Types and Names of Cyanide Complexes
Because there is such a variety of cyanide complexes, it is often difficult to compare
results of toxicological and environmental investigations of cyanide. For this reason
it is important that people responsible for managing cyanide understand and specify
the type of cyanide being measured. An essential chemical fact about cyanide is
that it is not an element like, for example, gold, arsenic and mercury. The free
cyanide ion comprises a nitrogen atom bonded to a carbon atom (CN-). This ion
combines with hydrogen to form hydrocyanic acid (HCN) and with metal ions to
form salts. The term cyanide is imprecisely applied to all of these forms and more
precise terms are defined as follows:
• Free cyanide-the sum of the free cyanide (CN-) ion and hydrocyanic acid,
HCN(aq). It is the free cyanide ion that is generally measured after suitable sample
treatment;
• Titratable cyanide-the cyanide concentration determined in solution by titration
with silver nitrate (AgNO3); often taken to mean free CN- but may include cyanide
from the dissociation of some cyano-metal complexes;
• simple cyanometal complexes-these contain only one type of metal ion, commonly
an alkaline or alkaline earth metal ion, and dissociate when dissolved in water to
release free cyanide;
• Complex cyanides-these contain more than one type of metal ion and dissociate in
water to release a metal ion and a cyanide-metal ion complex via a reaction of the
type, for example;
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where the complex ion may then dissociate further to give free cyanide;
• Total cyanide-the sum of all of the different forms of cyanide present in a system.
'Total cyanide' is a toxicologically meaningless term since its measurement requires
harsh sample treatment to break down intractable complex cyanides before free
cyanide can be measured;
• Weak acid dissociable (WAD) cyanide-cyanide that is readily released from
cyanide-containing complexes when the pH is lowered. Any free cyanide already
present and cyanide released from nickel, zinc, copper and cadmium complexes (but
not iron or cobalt complexes) is measured. WAD cyanide is measured by treating the
sample with a weak acid buffer solution such as a sodium acetate/acetic acid mixture
at pH 4.5 to 6. This is less harsh than the methods used for total cyanide. WAD
cyanide is generally considered to be the best current measure for assessing human
and animal toxicity;
• Cyanide amenable to chlorination (CATC)-an analytical quantity that requires
similar sample treatment to WAD but is much less reliable; and
• WAD CN, Available Cyanide and CATC generally measure the free and weakly
dissociable cyano-metal complexes
Different forms of cyanide can also be defined by reference to the various analytical
techniques and the types of cyanide that each technique is able to measure. For
example, free cyanide plus weak dissociable cyano-metal complexes may defined as
'WAD CN 4500-CN-I', 'CATC' or 'Available CN Method OIA-1677' depending on
the method used (adapted from Schulz, 2002). This approach to cyanide
nomenclature requires full disclosure of the method and the analytical protocol used
to ensure that meaningful comparisons and check analyses can be undertaken gas
2. Recycling or Disposing of Cyanide
As discussed in Section 4.3, even with the best recycling efforts, there may be
cyanide waste that needs to be treated to enhance its rate of degradation.
4. Natural degradation by oxidation
Reactions that cause the oxidative loss of the cyanide ion from aqueous alkaline
solutions are described by the following equations:
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Reactions 9 and 10 represent oxidative loss of cyanide ion giving rise to cyanate ion
and cyanogen gas respectively. As can be seen from the pH versus the
oxidation-reduction potential stability (Pourbaix) diagram presented in Figure 3, the
cyanate ion is thermodynamically the most stable form of cyanide under the majority
of operating and ambient environmental conditions. However, reaction 9 proceeds
very slowly in the absence of a catalyst so that loss of cyanide by this route is limited
during the ore extraction process.
Equation 8, in which cyanide is degraded by combining in aqueous solutions to
produce ammonia and bicarbonate, is particularly relevant to cyanide analysis and is
discussed further in Section 2.3 below
5. Cyanide: Health & Hygiene In The Workplace
6. Advances in the cyanidation of gold
In the last 15 years, most of the development in the cyanidation of gold has occurred
in response to the decreasing grade of deposits, the shift from surface mining to
underground mining, the increasing complexity of treatment and the concern for
environmental constraints. Research has focused on the optimization of reagent
addition (e.g. cyanide, oxygen and lead nitrate) and on metallurgical strategies to
measure and control these parameters, as well as in the areas of equipment and
automation. Efforts to introduce online cyanide analysers were made in the late
1980s; however, progress was slow.
The analyzers required additional testing to be reliable and effective and to yield
more accurate titration results. Online dissolved oxygen sensors were integrated in
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the late 1980s simultaneously with oxygen addition and various injection devices.
Sulfide in gold ores not only consumes oxygen and cyanide but also forms a coating
on gold grains. This passive layer reduces gold-leaching kinetics and overall
extraction. Improvement of lead nitrate addition strategy has made it possible to
minimize gold passivation.
7. Mechanism Of Cyanidation
Chemistry and electrochemistry
The most commonly used equation for the dissolution of gold in a cyanide solution,
known as Elsner‘s equation, is Gold dissolution is an electrochemical reaction in
which oxygen takes up electrons at one part of the metallic surface (the cathodic
zone), while the metal gives them up at another (the anodic zone).
According to this reaction, at low cyanide concentrations, the dissolution rate is a
function of cyanide concentration. At high cyanide concentrations, the dissolution
rate is a function of oxygen concentration. Gold dissolution is a heterogeneous
reaction that is controlled by the diffusion of both reacting species (O2 and CN_)
through the Nernst boundary layer. The rate of metal dissolution increases linearly
with increasing cyanide concentration until a maximum dissolution is reached,
beyond which there is a slight retarding effect. The dissolution rate is normally
mass-transportcontrolled in cyanide solutions with an activation energy of 8–20
kJ/mol. The formation of precipitates at the surface of gold grains is an important
aspect that determines the shape of the leaching kinetics plot.
8. Cyanide Management Plan
A key component in relation to treating cyanide is development of a sitewide
cyanide-management plan. The importance of properly developing and
administering such a plan has been highlighted by incidents at mine sites involving
the inadvertent release of cyanide to the environment. Aside from the potential for
environmental impact, such incidents broadly and negatively affect the image of the
mining industry and have led to emotional and damaging political responses, such as
the banning of cyanide in some regions.
Numerous guidance documents have been developed with regard to cyanide
management, and these documents should serve as a template for developing
site-specific cyanide management plans at mine sites.
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Implementation and adherence to a cyanide-management plan, augmented by
experienced scientific and engineering judgement, will help reduce both the number
and severity of environmental incidents involving cyanide.
The management of water and cyanide are intimately related, and development of a
cyanide-management plan should proceed in concert with development of a
water-management plan. A good cyanide-management plan will include descriptions
of how cyanide-containing solutions and slurries will be handled, stored, contained
and monitored, and in many cases the plan will also include a description of
treatment plants used to remove cyanide from solutions or slurries. At sites where
natural cyanide attenuation is important, the cyanide-management plan should
address the specifics of predicting and monitoring the effectiveness of the
attenuation processes. Decommissioning and closure are important phases in the life
cycle of cyanide management and should be addressed in the cyanide-management
plan.
9. Activated Carbon
The majority of the activated carbon used for precious metal recovery is either
granular coconut-shell carbon or peat-based extruded carbon. Important
considerations when selecting an activated carbon for use in a CIP operation include
gold-loading kinetics (activity), gold-loading capacity, elution kinetics, level to
which gold can be eluted, strength and abrasion resistance, particle-size distribution
and wet density. The properties of a particular activated-carbon sample will have a
significant impact on most aspects of the gold-recovery operation, affecting
variables such as carbon inventory, residence times, gold losses, carbon losses and
elution conditions. Therefore, due consideration must be given to the physical and
chemical properties of the virgin carbon when selecting a particular brand for
precious-metal recovery.
10. Carbon in pulp
Carbon in Pulp (CIP) is a technique for recovery of gold which has been liberated
into a cyanide solution as part of the gold cyanidation process, a gold extraction
technique.
Activated carbon acts like a sponge to aurocyanide and other complex ions in
solution. Hard carbon particles (much larger than the ore particle sizes) can be mixed
with the ore and cyanide solution mixture. The gold cyanide complex is adsorbed
onto the carbon until it comes to an equilibrium with the gold in solution. Because
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the carbon particles are much larger than the ore particles, the coarse carbon can then
be separated from the slurry by screening using a wire mesh
Flotation of gold :
The application of flotation on a reasonable scale within the gold-mining industry
commenced in the early 1930s following the introduction of watersoluble flotation
collectors (specifically xanthates and dithiophosphate collectors) that allowed
differential flotation of sulfide minerals). Prior to that time, a few gold mines in
Canada, Australia and Korea built flotation plants as the first step in the treatment of
complex and refractory gold ores. Flotation collectors on these plants were oils that
generated bulk low-grade gold concentrates, which were difficult to filter and dry.
Pre- and post-Second World War and up until late 1960s, most of the flotation
activity in the gold industry took place in Canada. During this period, Canada was
recognized as the second largest gold producer and a sizeable amount of the gold
production came from the treatment by flotation of copper–gold ores, refractory gold
ores and complex gold ores. The demand for sulfuric acid initiated by the booming
uranium industry during the late 1960s provided the catalyst for the installation of
pyrite flotation plants on numerous gold mines in South Africa. After roasting the
pyrite flotation concentrate to generate sulfur dioxide for the sulfuric acid plant, the
remaining calcine was cyanide leached to remove additional liberated gold. The
worldwide gold boom in the 1980s and 1990s created new opportunities in
Australasia and the Americas for the exploitation of medium-sized refractory gold
deposits by flotation and further treatment of the concentrates by bacterial and
pressure leaching. Many copper flotation plants around the world, and particularly
those in the Americas, have enough gold in the ore to ensure that special attention is
given to maximize the recovery of gold into the copper concentrate. A
comprehensive list of these operations, with details of the flotation reagent regimes
and circuit configurations that were in existence during the 1980s, is provided by
Bassarear. Since that time there has been a significant increase in the availability of
selective flotation collectors for gold and these are now widely in use on many large
and new copper flotation plants around the world.
4.3.1. Mineralogy
Gold occurs in a number of minerals (Harris, 1990) and the most important of these
is metallic gold and the gold metal alloys. Gold tellurides and gold–silver tellurides
are of less importance, although in particular deposits, they can account for a
significant proportion of the gold content. Aurostibite [AuSb2] and maldonite
[Au2Bi] are rare minerals but are present in some gold deposits. Aurocupride
[AuCu] is found in some primary copper ores. Gold particles in an ore deposit will
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vary in size from large nuggets to particles locked in the crystal lattice of sulfide
minerals. These sulfide minerals are referred to as gold carrier minerals and contain
trace to minor amounts of gold Often it is found that gold ores are refractory due to
the small size of the gold particles in the sulfides and concentration by flotation is
required, followed either by roasting, bacterial leaching or pressure leaching to
liberate the gold prior to cyanidation.
General aspects of gold flotation
Most of the reported fundamental work on the flotation of gold has been conducted
using high-purity gold and gold–silver alloys with the purpose of determining
collector–gold interactions and the nature of adsorption of collector ions or
molecules onto the gold surface. In addition, some work has been conducted to
decide whether or not pure gold has a natural hydrophobicity and hence some degree
of natural floatability. Flotation research work has been conducted on naturally
occurring native gold particles recovered from placer deposits and on gold particles
selected from lode deposits. The flotation characteristics of gold or gold minerals
found in refractory sulfide and copper ores have not been described in detail in the
literature. The sparse distribution of discrete gold minerals and particles, as well as
their exceedingly low concentration in ores, are the principal reasons for the lack of
fundamental work on gold flotation. A great deal of work has been reported on
specific ores, but such studies rarely distinguish between the flotation of native gold
and other gold minerals. Flotation of gold ores covers a broad field and it is a rather
difficult subject to generalize on. Most problems in gold ore flotation are not
connected with floating metallic gold. The flotation recovery of free gold
(throughout the text free gold is synonymous with liberated gold) is largely affected
by physical constraints such as the shape and size of the gold particles and the
stability of the froth. It is a generally accepted fact that liberated gold finer than about
150 mm floats readily with most collectors and in particular xanthates and
dithiophosphates. When free gold is floated with other sulfide minerals the extent of
bubble loading of sulfide particles may provide a barrier towards the attachment of
free gold, thereby reducing flotation performance. Research investigations have until
recently typically focused on the individual flotation behaviour of each gold-bearing
mineral in synthetic mixtures and not on mixtures of sulfide minerals in ‗real‘ ores.
In the flotation process, the main chemical effects are reagent type and pulp pH.
Recently, there has been a need to operate circuits at moderate pH levels, to improve
separation efficiencies when treating complex low-grade ores, to reduce costs of
reagents, to develop reagents that are stable over a wide pH range and to take
advantage of the synergistic benefit of mixture collector systems. This has led to
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ever-increasing research to develop new collectors and mixtures for the flotation of
gold-bearing.
Surface characteristics of pure gold
Several research papers have dealt with the hydrophilic and hydrophobic nature of
pure metallic gold and whether either a zero or a finite contact angle is observed.
Some workers measured contact angles on gold; while others did not. The
discrepancy between the findings of zero and of high contact angles on a pure gold
surface appeared to be due to both the presence of residual polishing agents and
organic contaminants. It is now generally accepted that the surface of pure clean gold
is hydrophilic and displays a zero contact angle. The hydrophilicity of gold is due to
the high Hamaker constant and is a result of the strong dispersion attraction forces
for water.
Collector less Flotation of Naturally Occurring
GOLD
Numerous references are to be found in the literature on the observed skin flotation
of gold, especially during the recovery of gold by gravity separation .On this
evidence gold was presumed to be naturally hydrophobic. As discussed above,
naturally occurring native gold surfaces are usually found to be hydrophobic; this is a
result of the contamination of the gold surface by organic compounds. Untarnished
gold of the appropriate particle size has been found to readily float with only a . The
earliest recorded laboratory work on gold flotation found that gold floated in the
presence of frother only, but not if its particle size was too large or if reagents such as
calcium oxide or sodium sulfide were added to the pulp. Gold can also be rendered
hydrophobic by the deposition of sulfur on the surface.
4.3.5.Collectors in Gold Flotation
Collector flotation of naturally occurring, placer and liberated gold
Gold hydrophobicity is enhanced by the addition of flotation collectors and no
flotation plant relies solely on the natural floatability of gold for its recovery.
Naturally occurring or free (liberated) gold is optimally recovered in a flotation
circuit at natural or near-natural pulp pH values and with the addition of small
amounts of collector. Inherently, naturally floating minerals float fast kinetically
Flotation tests on placer gold .showed that fine placer gold typically floated readily
with common sulfhydryl collectors and common frothers at natural pH without the
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addition of any special regulating reagents. Gold flotation recoveries ranged from 78
to 99%.
Flotation collectors for gold and gold carriers
Flotation with xanthate collectors involves the anodic oxidation of the collector that
may involve sub-processes such as metal xanthate formation, chemisorption of the
xanthate ion and oxidation of the xanthate to form dixanthogen. These adsorb onto
mineral surfaces, rendering the mineral hydrophobic. It is generally accepted that the
xanthate species responsible for the flotation of free gold is dixanthogen. This is a
neutral oil that will adsorb onto the surface of any naturally hydrophobic solid,
rendering it floatable. Dixanthogen may form on gold by either the application of an
applied potential or by a mixed potential mechanism in a pulp that involves the
reduction of oxygen. Studies have shown that the development of a finite contact
angle and the onset of flotation of gold particles occur at a potential close to that of
dixanthogen formation. The longer-chain xanthates are more readily oxidized,
generating dixanthogen at lower potentials. An increase in thiol chain length
increases the maximum contact angle, thereby increasing the hydrophobicity of the
surface species. Both these attributes favour the use of longer chain xanthates, such
as potassium amyl xanthate (PAX) for the flotation of free gold.
It is quite common to encounter silver and other precious metals forming alloys with
native gold. The positive effect that silver has on gold floatability was first
recognized in experiments using plates of pure gold, silver and gold–silver alloys.
The adsorption of ethyl xanthate on silver is generally thought to take place through
an electrochemical mechanism of metal xanthate formation on the surfaces . For
ethyl xanthate, the presence of silver in gold leads to silver xanthate formation at a
potential proportionately lower than for dixanthogen formation on pure gold . As a
consequence, the flotation of gold–silver alloys can be achieved at potentials
considerably lower than that for gold. Xanthate ions chemisorb on silver at potentials
below the region at which silver xanthate deposits. Chemisorbed ethyl xanthate
results in finite contact angles on silver surfaces and the initiation of flotation
appears to result from the chemisorptions process. For more rapid flotation
dixanthogen may play a supporting role. The chemisorbed sub-monolayer could be
important in retaining the dixanthogen at the gold surface through hydrophobic
interactions between the adsorbate and the bulk phase. The xanthogen formates are
produced by reacting alkyl chloroformate with xanthate salts. They are stable in
acidic conditions unlike the xanthates from which they are formed and are stable in
the pH range of 5–10.5. The formates appear to have superior pyrite rejection
properties compared to xanthates and dithiophoshate .
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Frothers in Gold Flotation
The strength and stability of the froth is important when floating free gold. There
appears to be a preference for polyglycol ether-based frothers on most gold plants in
combination with one or other frothers. When selectivity is required or, in the case of
copper–gold ores, where a copper concentrate is sold to a smelter, a weaker frother
such as methyl isobutyl carbinol (MIBC) is preferred. The choice of a particle
size-balanced frother is also an important consideration in gold flotation as this
promotes composite particle recovery in the scavenger flotation circuit. As a rule, the
glycol or polypropylene glycol methyl ether frothers are ideal for this application.
The blended interfroth frothers have found wide acceptance on Australian gold
plants and the base reagent is an alkyl aryl ester.
4.3.8.Activators in Gold Flotation
Activation implies improved floatability of a mineral after the addition of a soluble
base metal salt or sulfidizer. It is generally thought that the metal or sulfide ion
adsorbs onto the mineral surface thus changing its surface chemical properties. In
this way, the flotation response can be improved and/or the pH range of flotation for
the mineral can be extended, the rates of flotation increased and selectivity
improved.
Sulfidization
The application of sulfidizers (sodium sulfide and sodium hydrosulfide) to enhance
the flotation of oxidized ores is well. The first detailed laboratory study of the
influence of sodium sulfide on the flotation of gold-bearing ores was undertaken in
the mid-1930s. The outcome from this study was that, in general, sodium sulfide
retards the flotation of gold, although for some ores there was benefit in its addition.
Similar comments are to be found in the literature since that time. Sulfide ions appear
to act as flotation activators at low concentrations (less than 10_5M) and as a strong
depressant at concentrations above 10_5 M. The addition of sulfide ions converts
some coatings on mineral surfaces in sulfides and subsequent xanthate addition will
promote flotation. For successful activation, the sulfide activator should be added
slowly and at starvation quantities.
A recent study of a number of flotation plants found that floated gold grains in the
concentrate had a greater concentration of silver and sulfur than the gold remaining
in the tailings, implying that they assisted gold flotation. Laboratory flotation tests on
a flotation-plant tailing sample using NaHS and Silver ions recovered 30–45% of the
unfloated gold and NaSH provided the best results. Application of this information at
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plant scale resulted in a 7% increase in gold recovery at the Los Pelambres Mine in
Chile. Rejected gold particles at this mine were found to be coated with lead
carbonate.
A recent innovation, the Controlled Potential Sulfidization (CPS) process, where the
pulp potential is controlled has been successfully applied at a number of copper–gold
ore flotation plants in Australia In this process, sodium sulfide or sodium hydrogen
sulfide is added and the solution potential is controlled at about _450mV prior to
flotation.
4.3.9. Depression of Gold In Flotation
Depressants for native gold that are usually introduced during the flotation process
include compounds such as calcium ions, chloride ions, calcium carbonate, cyanide,
sodium silicate, sodium sulfite, ferric and heavy metal ions, tannin and related
compounds, starch and other organic depressants and many others . All of these may
competitively adsorb on the gold surface thus preventing the adsorption of the
collector(s) added. It has also been suggested that the ferric ions, which would be in
the form of hydrated oxides, may act as a physical barrier between the air bubble and
gold surface but this effect is reversed simply by washing with water However,
flotation of native gold often proceeds satisfactorily in the presence of many of these
compounds. In general, the results reported by different authors are not in good
agreement (Allan and Woodcock, 2001). It is likely that other components in
solution or on the surface of the gold that were not measured provide the answer for
the different outcomes. Lime cannot be considered as just a pH modifier and studies
have shown that calcium is strongly adsorbed on sulfide minerals and gold at pH
values at and above 10. This adsorption is enhanced if excess sulfate in the pulp
promotes calcium-sulfate coatings on particles.
Desorption of calcium from the surface by reducing the pH can be assisted by the use
of specific calcium-complexing ions such as polyphosphate. Furthermore, if the
calcium release is attempted while adding excess activator, then a hydrophilic
hydroxide coating can result. Metals ions introduced from the circuit water, or from
soluble metal ions in the ore, may adsorb and nucleate as hydroxide coatings on all
particle surfaces, thus inhibiting collector adsorption. The recommended method of
flotation treatment is to operate at as low a pH value as practical, avoid rapid
increases in pH, add activator slowly or condition separately and keep the tailings
dam at a pH of minimum solubility.
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4.3.10 Flotation of Gold and Gold-Bearing Minerals
Differential flotation of natural and liberated gold
Experience has shown that free gold particles can be recovered selectively against
pyrite, by keeping the gold particle surfaces as clean as possible of organic species
and by removing any adhering slime particles This can be achieved with the use of
little or no pH regulators, only small dosages of collectors and suitable frother to
stabilize the froth, and possibly a small amount of dispersant. Selectivity for gold
against pyrite was found to be enhanced in the presence of collectors such as alkoxy
or phenoxy carbonyl alkyl thionocarbamates, dialkyl or diaryl monothiophosphates
and monothiophosphinates, glyoxalidine and aminothiophenols.
Monothiophosphorous acids have been shown to be able to float gold
Selectively from base metal sulfides.
The use of hydrogen peroxide as an oxidizing agent in the selective flotation of gold
from pyrite with PAX has been demonstrated at laboratory scale. Hydrogen peroxide
addition by itself rendered both the gold and pyrite surfaces hydrophobic. The
addition of xanthate converted the gold surface into a fairly hydrophobic condition,
whereas the pyrite was still hydrophilic at pH values of 10 and higher. On most
flotation plants, there is a tendency to treat sulfide ores containing free gold as
though the gold is associated in a massive or complex sulfide mineral matrix. This
leads to high dosage levels of collector and activator addition. In this application,
xanthate adsorption on both sulfides and gold makes selective flotation rather
difficult due to the formation of dixanthogen on both the gold and sulfide surfaces.
4.3.12. Flotation of gold-carrying iron sulfides
Pyrite is known to be oleophilic when the surface is free of oxidation products but it
is nevertheless necessary to use a collector to float pyrite. Thiol collectors are the
most commonly used Pyrite can be recovered optimally in either acidic or alkaline
conditions. Xanthates are the primary collectors on the alkaline side and MBT is
used when floating in an acidic medium. Blends of different collectors are
particularly effective for recovering pyrite and mixtures of xanthate, thiocarbamates
and xanthate and xanthate mixed with MBTs have been successfully used.
Dithiophosphates are usually used as secondary collectors and on their own are
reported to be selective against pyrite and to a lesser extent against arsenopyrite.
Aminebased collectors are capable of floating cyanide-leached pyrite without the use
of acid pre-conditioning.
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Arsenopyrite has very similar properties to pyrite and the flotation conditions for its
recovery are similar to pyrite. Arsenopyrite is susceptible to the formation of
oxidation products on its surface. Under oxidizing conditions, ahigh concentration of
PAX is required in order to achieve high arsenopyrite recovery. The reason for this is
not clear but the low rate of arsenopyrite floatability could be attributed to the
formation of iron-oxide species on the surface of the arsenopyrite. Oxidation
products such as ferric and ferrous ions are also present in the pulp and depending on
the chemical conditions these promote the formation of dixanthogen and bulk
ferric-xanthate compounds in solution. Pyrrhotite floats readily in acid and neutral
pH ranges. Surface coatings in the alkaline regions may result in depressed flotation
recovery, while collector regimes are similar to those for pyrite flotation. Pyrrhotite
oxidizes readily and this makes it more difficult to float than arsenopyrite. Oxidation
products in solution create flotation problems similar to those mentioned above for
arsenopyrite. The floatability of marcasite appears to be variable as it has been
shown to float more readily than pyrite, while more recently.
4.3.14. Flotation of copper– gold ores
For the bulk flotation of copper minerals and gold from supergene copper ores, it is
normal practice to add a xanthate as the primary collector and dithiophosphate as a
secondary collector. This combination gives satisfactory results in respect of copper
concentrate grade and copper and gold recoveries. When pyrite is present in the ore,
the choice of collectors will generally depend on the ratio of copper to pyrite in the
feed. Dithiophosphate collectors are found to be more selective against pyrite and
better gold collectors than xanthate . With low pyrite contents in the feed, the
ultimate gold recovery is dependent on the type of xanthate selected, longer
carbon-chain lengths achieving higher gold recovery. Ores containing large amounts
of pyrite require collectors that are selective towards both the copper minerals and
free gold. For these ores, the preferred collectors are the dithiophosphates and the
new generation of ‗gold‘ collectors. Recent laboratory testwork on a copper–gold ore
containing pyrite showed that Aerofloat 7249 achieved the highest gold . Aerofloat
6697 provided the best gold concentrate grade at pH values less than 11.5 while
Aerofloat 7249 and Aerofloat 208 were better at pH values above 11.5 and gold
grades in excess of 250 g/t were generated. At industrial scale, the use of Aerofloat
7249 provided a 4.5% improved gold recovery at the Freeport Copper Mine, while
the addition of 3477 in the cleaner scavenger circuit improved the recovery of
tarnished flaky gold at the Kemess Copper–Gold Mine in Canada.
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4.3.15. Particle size and shape in flotation
It is well known that particle size is an important parameter in flotation and that size
limits exist at which minerals will and will not float. The high particle-density of
gold and its malleable and ductile properties that favour the propagation of platy
particles, further compound this effect. Platy/flaky particles are formed in the
treatment process, particularly in grinding, or during transportation events in nature .
During these events, some gold particles are impregnated with nonfloatable particles
(Taggart, 1945; Pevzner et al., 1966), inhibiting flotation. Passivation of a
gold-particle surface may also occur after considerable hammering by steel
grinding-media . On the other hand, it is postulated that the surface of the gold could
become more active and therefore more floatable due to work hardening. It has been
suggested that the practical particle size limits for gold flotation are around 5–200
mm. Particles as small as 3 mm have been floated at laboratory scale), while actual
measurements indicate that the flotation performance on many gold plants decrease
rapidly below 10 mm (Chryssoulis, 2004). At the coarse end, gold particles as large
as 300 mm (Leaver and Woolf, 1934b) and 700 mm have been floated in laboratory
flotation cells under specific operating conditions and high collector additions (Lins
and Adamian, 1993). Flotation of 590 mm gold particles has been reported on an
industrial scale with ‗unit‘ flotation cells (Leaver and Woolf, 1934). Pulp density and
aeration rates influence flotation-cell pulp hydrodynamics and are important
parameters in extending the particle-size limits of gold flotation. There is conflicting
commentary on the best pulp density for gold particle flotation, both a high
pulp-density (Leaver and Woolf, 1934a) and a low pulp-density being
recommended.
4.3.16. FLOTATION CIRCUITS
Flotation circuit configuration on most gold mines can be divided into a number of
categories, viz. open circuits with no cleaning at all, and open and closed circuits
with single stage and two stages of cleaning. Open circuits have the advantage of no
feedback from the effects of non-steady-state operation and therefore are inherently
more stable than the closed-circuit configuration.
Closed and open-circuit flotation cleaning is used on gold mines where high-grade
concentrates are required for roasting and smelting. Under these conditions, it is
difficult to maintain very high gold and sulfide flotation recoveries, while also
producing an acceptable grade of concentrate. Where there is no constraint on
concentrate quality, high gold and sulfide flotation recoveries are achievable to the
extent that a discardable gold flotation tail is possible. Cleaning-circuit
configuration, either single or two stages of cleaning, and cleaner residence time are
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related to the particle size of the sulfides in the flotation feed and also the presence or
absence of floatable gangue components. Unit flotation cells and the more recent
Flash flotation cells are installed in grinding circuits with the purpose of improving
the overall flotation recovery of free gold . The aim is to remove as much of the free
gold contained in the circulating load of the grinding mill before it is overground or
is covered with coatings of iron, sulfide or other coatings that will lower flotation
recoveries. Improved overall gold flotation recoveries of 2–10% have been quoted
(Sandstrom and Jonsson, 1988; Jennings and Traczyk, 1988; McCulloch, 1990).
Furthermore, the inclusion of Unit and Flash flotation cells will generally provide
better flotation stability and performance. Improved overall gold flotation recoveries
from 3% to 10%
Slime coatings and floatable non-sulfide gangue
The deleterious impact of clay slimes on gold flotation is well. The failure of free
gold and sulfide minerals to float has at times been shown to be related to the
presence of coatings of colloidal or near-colloidal gangue or silicate material
adhering to the mineral surface.
These coatings are formed under pulp conditions in which the sulfide particles and
silicate particles are oppositely charged. Gangue minerals that are known to cause
problems include talcose and carbonaceous minerals, bentonite clay, goethite
[FeO(OH)] iron oxide and manganese slimes, pyrophyllite [AlSi2O5OH] and
carbonates . Slime coatings are controlled by the use of gangue-dispersing agents.
Sodium silicate is widely used for this purpose and is most effective when the
alkalinity is carefully controlled. Sodium sulfide has also been found to be an
effective dispersing agent. In addition to coating the mineral surface, the gangue
particles may coat the bubble surface, affecting the ability of any gold and sulfide
particles to attach to the air bubbles. Other more recent remedies to overcome the
problem of slime coatings have included physical methods such as removal of the
slimes by.
Organic compounds of high molecular weight that maintain a state of dispersion of
deleterious slime components by forming wettable coatings on the gangue particles
are used for much the same purpose as the inorganic dispersing agents. These
organic compounds are referred to as organic gangue depressants. Typical examples
are glue, starch, dextrin, gum arabic, carboxymethylcellulose and the more recent
modified-guar gums. Selection of the correct depressant type and dosage is critical,
as an overdose results in both loss of free gold and sulfides that contain gold. The
anionic polymers generally have a negligible depressant capability on sulfide
minerals while the cationic polymers are capable of acting as sulfidemineral
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depressants .The combination of a collector and depressant is also important since in
the flotation of pyrite, for example, guar gum will have a more adverse effect when
used with MBT than with xanthate . An alternative approach that has had some
success is to add small quantities of frother to the pulp and selectively float the
talcaceous minerals prior to removing the bulk sulfide concentrate. These talc
concentrates may contain up to 30%–40% of the gold contained in the feed and this
concentrate can either be cyanide leached separately or recombined with the sulfide
tailing prior to cyanidation .
Porphyry copper–gold ores usually contain some gangue components that are highly
floatable and contaminate the copper concentrate. Maintenance of a high
copper-concentrate grade requires that gangue depressants be used. Silicates, guards
and carboxymethylcellulose are the common depressants applied in the copper
industry.
Carbonaceous and graphitic minerals are soft and flaky, and easily broken down
during grinding. During flotation, the carbon floats readily owing to its fine grain
size, natural hydrophobicity, platy nature and low density. Graphitic carbon and
clays can be the cause of poor gold recovery on many refractory gold or flotation
plants.
Natural metal and organic coatings on gold
Most coatings on mineral surfaces are detrimental to flotation, but in some cases, the
effects can be overcome. Many types of surface coatings have been reported to
occur on native gold particles. Perhaps the most difficult coatings to cope with are
hydrated iron oxides. The surfaces of gold particles can become coated naturally
with precipitates of iron, from oxidized sulfides in an orebody or from rusting iron,
such as iron grinding media, as first reported by Head (1936). Gold from placer
deposits heavily stained or coated by iron oxides or impregnated by hydrophilic
minerals is not easily floated. Tarnished gold has been found also to have a markedly
higher mercury content compared to ‗shiny‘ gold .Gold particles coated with
manganese dioxide have also been reported . Some gold flotation pulps may contain
humic and tannin substances) and sulfide ions from sulfide mineral that are reported
to impact on gold flotation. Humic acid has been found to be only marginally
deleterious to gold flotation and some naturally occurring organic coatings can be
removed by conditioning with sodium hydroxide or acid solution .
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Factors For Consideration In Refractory Process Selection
Based on GRD Minproc‘s experience on numerous projects, the following factors
are considered to be of importance in selecting a process for treatment of refractory
gold ores. All factors should be taken into account at an early stage and process
options should be kept open for as long as possible due to the potential for
unforeseen issues to impact on the economics of a particular process.
Refractory gold ores
Successful concentration of gold in refractory sulfide ores is almost exclusively
dependent on the association of the gold with the sulfides . Refractory gold ores
commonly contain free gold, sub-microscopic gold, carbonaceous material, base
metals, pyrite, marcasite, arsenopyrite and pyrrhotite . Clays and graphitic carbon are
the most troublesome accessory components in some of these ores, as far as gold
concentration is concerned. Arsenopyrite has very similar properties to pyrite and the
flotation conditions for its recovery are similar to pyrite. Arsenopyrite is marginally
less hard and more brittle than pyrite and pyrrhotite . During milling, the
arsenopyrite is therefore subject to more overgrinding than pyrite and pyrrhotite. The
difference in the recovery of these minerals is due not only to the difference in the
surface chemical properties of the particle but also to the difference in their overall
size distribution
.
Reference
Advances in gold ore processing
A method for leaching or dissolving gold from ore
environmental management in mining