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Designing a process for the recovery of Nickel-Copper alloy from Sirosmelt slag at Empress NickelRefinery

D. Muchena 2003 1

1.0 I NTRODUCTI ON

Rio Tinto Zimbabwe Limited (RioZim) operates the Empress Nickel Refinery

(E.N.R) at Eiffel Flats, Kadoma, Zimbabwe [1]. It was built in 1968 to refine

nickel-copper matte from the Empress nickel mine using the Outokumpu

atmospheric leaching process. The mine and its associated concentrator and

smelter were closed down in 1982 after the economic mineral reserves had

been depleted. The refinery has been operating successfully as a toll refinery

since then, treating mainly matte from the Bamangwato Concession Limited

(B.C.L) operations in Botswana. This matte contains about 45-50% nickel, 40-

45% copper and 5-7% sulphur with the remainder made up of cobalt, iron and

arsenic. The refinery aims to maximize copper and nickel recoveries and to

produce nickel and copper cathodes that are satisfactory to the customers

requirements in terms of quality, quantity and size.

Sirosmelt slags are produced during the smelting of residues from

thickener 4 from the cementation and copper 1 leach circuit at E.N.R., the

slags being molten by-products of high temperature processes that are

primarily used to separate the metallic and nonmetallic constituents

contained in the bulk residues. Residue smelting is a process in which one

of the crucial roles slag plays, is the removal of impurities from the matte.

After a heat of the smelting process upon tapping, two different layers of

the slag and matte should ideally be formed due to differences in densities

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of the slag and matte, whereupon the matte is tapped onto the granulation

pit and the slag tapped into chill moulds. The matte being tapped should

ideally be free of any slag and the slag should also ideally be free of any

matte. However due to conditions prevailing in the furnace during any

heat, some of the matte gets entrained in the slag, while some matte is

oxidised due to the highly oxidising conditions that prevail in the furnace.

The oxidised matte reports to the slag. These occurrences are referred to

as metal-to-slag losses. It is envisaged that the total metal-to-slag losses

(i.e. total of Ni + Cu) should not exceed 1% upon assaying any sample of

Sirosmelt slag. However since the commissioning of the furnace high

metal-to-slag losses have been experienced, exceeding 1%. This has lead

to a lot of metal being lost with the slag being dumped at the slag dump at

E.N.R. To date the total of metal lost exceeds 250 tons (see appendix 1),

which is a loss in revenue exceeding U.S.$1,176,696,72. Thus against this

background the aim of this project is to design a suitable and cost effective

process to recover the Cu-Ni alloy from Sirosmelt slag at E.N.R.

The method considered for this project involves mineral processing since

it is a relatively inexpensive process with very little complications in the

process. Effective recovery of the alloy from slag involves separation of

the alloy from slag to acceptable recoveries. This can be achieved through

considering the differences in physicochemical properties between the

values and the gangue such as specific gravity, size, shape, colour, and

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electrical and magnetic properties. Such differences can then be

manipulated in the design of a suitable process for the physical separation

of the alloy from the slag. In order to determine such physicochemical

properties as mentioned above, characterization of the slag is necessary.

This involves a series of mineralogical analysis of the slag, such as

chemical analysis, sieve tests and metallographic analysis. Information

from the mineralogical analysis will then be used to design a preliminary

experimental flow sheet for laboratory scale, milling and physical

separation investigations to predict the behavior of the slag to such

techniques. Laboratory scale separation will be carried out. Results from

these tests will lead to design of a plant/pilot plant scale flow, diagram for

the milling of, and separation of the alloy from the slag.

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2. 0 LI TERATURE SURVEY

2. 1 Slag Characterizat ion

Slag characterisation involves investigations leading to an understanding of the

nature of the entrained alloy particles in the slag and their association with other

components in the slag [2]. This when accomplished can help in establishing the

slag to metal/alloy proportions, mode of occurrence of the entrained copper and

nickel and compositional variations within the slag stockpile. The dissemination

of the entrapped metal of interest within the slag matrix has a strong bearing on

the process technology to be used. Information from the study of slag

characteristics would help in predicting the degree of grinding required for

effective liberation of values from the gangue and effectiveness of the separation

methods in concentrating the Cu-Ni alloy.

The first stage in the characterization of the slag would involve chemical analysis

of the slag samples. Chemical analysis at Empress Nickel Refinery is carried out

for every slag-tap conducted after every heat of the Sirosmelt furnace. As the

slag is being tapped into chill moulds, a sample is taken and sent for analysis to

assay for different elements and other components in the slag and their relative

abundances. X-ray analysis is used in performing chemical analysis on the slag.

This is a non-destructive technique for confirming a samples elemental

composition. X-rays are bombarded on the sample, and a detector measures the

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secondary X-rays emitted from the sample. From these measurements a

computer can calculate a quantitative analysis of the sample.

Particle size analyses can also be carried out on slag samples. These are a

range of tests performed with the aim of determining the amount of crushing to

be performed on the slag in order to find an optimum degree of liberation of the

metal values from the gangue. These include sieve analysis and recovery versus

particle size analysis on the separation equipment of choice [3]. Sieve analyses

provide information on the size fractions to which the greatest proportion of metal

values report, leading to predictions of the optimum mesh of grind for effective

liberation of metal values from the gangue. A metal distribution curve will be

plotted for various size fractions to which the slag would have been ground [4].

Recovery versus particle size analysis shows the particle size for which the

greatest recovery is achieved for a particular particle separation process e.g.

gravity separation.

A slag characterization exercise would thus provide information on the degree of

comminution needed and the beneficiation methods necessary. It also shows the

nature of the final products to be obtained and what further processing is

necessary.

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2. 2 Part icle Size Analysis

Size analysis of the various products of a mill constitutes a fundamental part of

laboratory testing procedure [5]. In this project, it will be of great importance in

determining the quality of grinding and in establishing the degree of liberation of

the value metal alloy from the gangue at various particle sizes. In the separation

stage, size analysis of the products will be used to determine the optimum size of

the feed to the process for maximum efficiency and to determine the size range

at which any losses are occurring in the process, so that they may be reduced

[3].

The primary function of precision particle size analysis is to obtain quantitative

data about the size distribution of particles in the material [5]. However, the exact

size of an irregular particle cannot be measured. For a spherical particle, the size

is uniquely defined by its diameter, while for irregular particles the equivalent

diameter is often used. Recorded data from any size analysis should, where

possible, be accompanied by some remarks, which indicate the approximate

shape of the particles. Descriptions such as granular or acicular are usually

quite adequate to convey the approximate shape of the particle in question. A

short list of some of the more common methods of size analysis, together with

their effective size ranges is given in table 2.1 [5].

Test sieving is the most widely used method for particle size analysis. It covers a

very wide range of particle size; and will thus be used in this project for any

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particle size investigations. Sieve analysis is accomplished by passing a known

weight of sample material, successively through finer sieves, and weighing the

amount collected on each sieve to determine the percentage weight in each size

fraction. Sieving is carried out with wet or dry materials and the sieves are

usually agitated to expose all the particles to the openings.

Method Approximate useful range (microns)

Test Sieving 100000-10

Elutriation 40-5

Microscopy (optical) 50-0.25

Sedimentation (gravity) 40-1

Sedimentation (centrifugal) 5-0.05

Electron microscopy 1-0.005

Table 2.1: Some common methods of size analysis

In each of the standard series of sieves the apertures of consecutive sieves bear

a constant relationship to each other. It has been realized that a useful sieve

scale is one in which the ratio of the aperture widths of adjacent sieves is the

square root of two (i.e. 7KHDGYDQWDJHRIVXFKDVFDOHLVWKDWWKHDSHUWXUHareas double at each sieve, facilitating graphical presentation of results.

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There are several ways in which the results of a sieve test can be tabulated. The

methods (usually tabular) should show [3]:

1 The sieve size ranges used in the test.

2 The weight of material in each size range.

3 The weight of material in each size range expressed as a percentage of

the total weight.

4 The nominal aperture sizes of the sieves used in the test.

5 The cumulative percentage of material passing through the sieves.

6 The cumulative percentage of material retained on the sieves.

The results of a sieving test should always be plotted graphically in order to

assess their full significance. There are many different ways of recording results,

the most common being that of plotting cumulative undersize (or oversize)

against particle size. Many curves of cumulative oversize or undersize against

particle size are S-shaped as in fig. 2.1.

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2. 3 Milling Operat ions

Milling is carried out in order to prepare the slag from the slag dump, for

extraction of the valuable metal [5]. Apart from regulating the size of the slag, it is

a process of physically separating the grains of valuable metal from the gangue

minerals, to produce an enriched portion, or concentrate, containing most of the

valuable minerals, and a discard, or tailing, containing predominantly the gangue

minerals. The first stage in a milling process is comminution, whose major

objective is the liberation, or release, of the valuable mineral, and in this case,

alloy, from the associated gangue minerals by size reduction of the slag size to

the coarsest possible particle size. Comminution in its earliest stages is carried

out in order to make the freshly excavated material easy to handle by scrappers,

conveyors, and slag carriers [5]. Comminution in the mill takes place as a

sequence of crushing or grinding processes. Crushing reduces the particle size

of run-of-dump slag to such a level that grinding can be effected until the metal

alloy and the gangue are substantially produced as separate particles. Crushing

is usually a dry process, and is performed in several stages with small reduction

ratiosranging from three to six in each stage. The reduction ratio in crushing can

be defined as the ratio of maximum particle size entering to maximum size

leaving the crusher.

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2. 3. 1 Crushing Equipment

Crushing is the first mechanical stage in the process of comminution of the slag

material in which the main objective is the liberation of the valuable alloy from the

gangue [5]. The nature of machinery for a given crushing operation is influenced

by the nature of the product required and the quantity and size of material to be

handled. Lumps of slag to be fed can be as large as 50cm and these have to be

reduced in the primary crushing stage to 10-20cm in heavy-duty machines.

Crushing is normally done in open or closed-circuit depending on product size. In

open-circuit crushing, undersize material from the screen is combined with the

crusher product and is routed to the next operation. If the crusher is producing

ball-mill feed it is good practice to use closed-circuit crushing in which the

undersize from the screen is the finished product. Closed-circuit crushing has the

advantage of giving greater flexibility to the crushing plant as a whole.

Crushers are classified as primary or secondary crushers. In the primary stage

crushing of slag lumps from the slag stockpile, heavy-duty machines are used to

reduce the slag lumps to a size suitable for transport by conveyors and for

feeding the secondary crushers. Two suitable primary crushers are the jaw and

gyratory crushers, as illustrated in figs. 2.2 and 2.3 [5]. Jaw crushers range in

size up to1680mm gape by 2130mm width. This size of machine can handle slag

lumps with a maximum size of 1.22m at a crushing rate of approximately 725t/h

with a 203mm set. However, at crushing rates above 545t/h the economic

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advantage of the jaw crusher over the gyratory diminishes; and above 725t/h jaw

crushers cannot compete with gyratory crushers.

Jaw crushers may be divided into three main groups; the Blake, with a movable

jaw pivoted at the top, giving greatest movement to the smallest lumps, the

Dodge, with the movable jaw pivoted at the bottom, giving greatest movement to

the largest lumps, and the overhead eccentric, which is hinged at the top similarly

to the Blake, with the movable jaw suspended on the eccentric shaft. Jaw

crushers are applied to the primary crushing of hard materials and are usually

followed by other types of crushers. In smaller sizes they are used as single-

stage machines.

Figure 2.2 Cross-section through single-toggle jaw crusher

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The gyratory crusher consists of a cone-shaped pestle oscillating within a larger

cone-shaped mortar or bowl. The angles of the cone are such that the width of

the passage decreases towards the bottom of the working faces. The pestle

consists of a mantle, which is free to turn on its spindle. The spindle is driven

from an eccentric bearing below. Differential motion causing attrition can occur

only when pieces are caught simultaneously at the top and bottom of the

passage owing to different radii at these points. Crushing occurs through the full

cycle in a gyratory crusher, and this produces a higher crushing capacity than a

similar sized jaw crusher, which crushes only in the shuttling half of the cycle. For

this reason gyratories are often operated in parallel with a scalping grizzly

screen, provided the added cost of the screen is less than the cost of increased

crusher capacity.

Fig 2.3 A functional diagram of a gyratory crusher

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Secondary crushers are much lighter than the heavy-duty, rugged primary

machines. Since they take the primary crushed ore as feed, the maximum feed

size of slag material it can take will be ideally less than 15cm in diameter.

Secondary crushers also operate with dry feeds, and their purpose can be to

reduce the slag material to a size suitable for grinding. The bulk of secondary

crushing of metalliferous ores is performed by cone crushers, although crushing

rolls and hammer mills are used for some applications.

Another class of crushers in which comminution is by impact rather than

compression, by sharp blows applied at high speed to free-falling rock, exists.

This class is called impact crushers [6]. The moving parts are beaters, which

transfer some of their kinetic energy to the ore particles on contacting them. The

internal stresses created in the particles are often large enough to cause them to

shatter. These forces are increased by causing the particles to impact upon an

anvil or breaker plate. Examples of impact crushers include the hammer mill, the

impact milland the Tidco Barmac Crusher[7].

The Tidco Barmac crusher combines impact crushing, high-intensity grinding and

multi-particle pulverizing, and as such, is best suited in the tertiary crushing or

primary grinding stage, producing products in the 0.06-12mm size range. A

cross-section of the Duopactor, which can handle feeds of up to 650t/h, at a top

size of over 50mm is shown in fig. 2.4. The basic comminution principle

employed involves acceleration of particles within a special ore-lined rotor

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revolving at a high speed. Grinding will commence when rock enters the rotor,

and is thrown centrifugally, achieving exit velocities of up to 90 metres per

second. The rotor continuously discharges into a highly turbulent particle cloud

contained within the crushing chamber, where reduction occurs primarily by

rock-on-rock impact, attrition and abrasion.

Fig 2.4 Cross-section of a Barmac Duopactor Crusher (Courtesy www.teara.govt.nz)

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2. 3. 2 Grinding Mills

Grinding is normally the last stage in the process of comminution. In this stage

the particles are reduced in size by a combination of impact and abrasion, in

suspension in water. It is performed in rotating cylindrical vessels known as

tumbling mills[5]. These contain a charge of loose crushing bodies-the grinding

medium-which is free to move inside the mill, thus comminuting the slag

particles.

It is the purpose of the grinding section to exercise close control of the product

size. Thus investigations need to be carried out first to determine an optimum

mesh of grind in order to exercise close control of the product from any slag

grinding section. The optimum mesh of grind will depend on many factors,

including the extent to which the values are dispersed in the gangue, and the

subsequent separation process. The use of a grinding mill in the comminution of

slag form E.N.R. is envisaged to be on tailings from the first stage of gravity

concentration MZT1 (see fig. 3.1). Grinding is to be performed wet, being a

continuous process with material being fed at a controlled rat from storage bins

into one end of the mill and overflowing at the other end after a suitable dwell

time.

Tumbling mills are of three basic types: rod, rod and autogenous [6]. At Empress

Nickel refinery, ball mills (see fig. 2.5) are extensively used on the cementation

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leach circuit and on the siroleach circuit [1]. Thus a ball mill would be suitable for

grinding purposes on crushed slag from secondary crushers since such a mill

would be readily available and provisions for its use on slag recovery can be

made. A ball mill uses steel balls as the grinding medium. Closed circuit grinding

with high circulating loads would be preferable to open circuit grinding since the

former produces a closely sized end product and a high output per unit volume

compared with open circuit grinding. Grinding in a ball mill is effected by point

contact of balls and slag particles and given time, a desired degree of fineness

can be achieved.

Figure 2.5 A ball mill (courtesywww.mine-engineer.com)

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2. 4 Classif icat ion

Classification is a method of separating mixtures of minerals into two or more

products on the basis of the velocity with which the grains fall through a fluid

medium [5]. This fluid is usually water, and wet classification is generally applied

to mineral particles, which are considered too fine to be sorted efficiently by

screening. Since the velocity of particles in fluid medium is dependent not only on

the size, but also on the specific gravity and shape of the particles, the principles

of classification are important in mineral separations utilizing gravity

concentrators. Classifiers consist essentially of a sortingcolumn in which a fluid

is rising at a uniform rate. Particles introduced into the sorting column either sink

or rise according to whether their terminal velocities are greater or less than the

upward velocity of the fluid. The sorting column therefore separates the feed into

two products - an overflowconsisting of particles with terminal velocities less

than the velocity of the fluid and an underflowconsisting of particles with terminal

velocities less than the velocity of the fluid and an underflowor spigot productof

particles with terminal velocities greater than the rising velocity.

Many different types of classifiers have been designed and built which include

the settling cone(fig. 2.6), which is sometimes used as a dewatering unit in

small-scale operations [8]. It is often used in the aggregate industry to deslime

coarse sands products.

The rake-classifierutilises rakes actuated by an eccentric motion, which causes

them to dip into the settled material and to move it up the incline to the discharge.

Spiral Classifiers(fig. 2.7) use a continuously revolving spiral to move the sands

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up the slope. They can be operated at steeper slopes than the rake classifiers, in

which the sands tend to slip back when the rakes are removed.

Figure 2.7 Spiral classifier

The hydrocyclone(fig. 2.8) is a continuously operating classifying device that

utilises centrifugal force to accelerate the settling rate of particles. It is one of the

most important devices in the minerals industry [9], its main use in mineral

processing being as a classifier, which has proved extremely efficient at fine

separation sizes. It is widely used in closed circuit grinding operations but has

found many other uses such as de-sliming, de-gritting, and thickening. In the

recovery of alloy from E.N.R. slag, classification can be used before gravity

separation. Since gravity separators are extremely sensitive to the presence of

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slimes, which increase viscosity of the slurry and hence reduce the sharpness of

separation, and obscure visual cut-off points [5]. Settling cones preceding the

gravity device can be used to control pulp density within the circuit. For a

substantial increase in pulp density, hydrocyclones or thickeners may be used.

Figure 2.8 Hydrocyclone

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2. 5 Gravity Concent rat ion

Gravity concentration methods separate minerals of different specific gravity by

their relative movement in response to gravity and one or more other forces, the

latter often being the resistance to motion offered by a viscous fluid such as

water or air [5]. It is essential for effective separation of alloy particles from

gangue, that a marked density difference exists between the minerals and the

gangue. The concentration criteriongives the relative effectiveness of gravity

separation methods on a particular ore type.

).(.).(.

).(.).(.

mFluidMediuGSeralLighterMinGS

eralLighterMinGSeralHeavierMinGSonionCriteriConcentrat

=

When the quotient is greater than 2.5, whether positive or negative, then gravity

separation is relatively easy, the efficiency of separation decreasing as the value

of the quotient decreases. The motion of a particle in a fluid is dependent not

only on its specific gravity, but also on its size. Large particles will be affected

more than smaller ones. The efficiency of gravity processes therefore increases

with particle size, and the particles should be sufficiently coarse. In practice,

close size control of feeds to gravity processes is required in order to reduce the

size effect and make the relative motion of the particles specific gravity

dependant.

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2.5.1 Gravity Separators

It is essential for the efficient operation of all gravity separators that the feed is

carefully prepared through grinding, successive regrinding of middlings,

desliming and screening. One of the most important aspects of gravity circuit

operations is correct water-balance within the plant. Almost all gravity

concentrators have an optimum feed pulp-density, and relatively little deviation

from this density causes a rapid decline in efficiency. Accurate pulp-density

control is therefore essential, and this is most important on the raw feed [5]. This

can be achieved through the use of settling cones and such other thickeners,

preceding the gravity device. A selective but comprehensive review of gravity

separators, which are best suited for purposes of recovering alloy form ground

slag, is given in the following sections.

2.5.1.1 Spirals

Spiral concentrators find many varied applications in mineral processing. A spiral

is composed of a helical conduit of modified semicircular cross-section. Feed

pulp of between 15-VROLGVE\ZHLJKWDQGLQWKHVL]HUDQJHPPWR PLV

introduced at the top of the spiral and, as it flows spirally downwards, the

particles stratify due to the combined effect of centrifugal force, the differential

settling rates of the particles and the effect of interstitial trickling through the

flowing particle bed. These mechanisms are complex being much influenced by

the slurry density and particle size. It is believed in some academic quarters that

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the main separation effect is due to hindered settling, with the largest densest

particles reporting preferentially to the concentrate, which forms in a band along

the inner edge of the stream (fig. 2.9) [10].

Ports for the removal of the higher specific gravity particles are located at the

lowest points in the cross-section. Wash-water added at the inner edge of the

stream, flows outwardly across the concentrate band. The width of the

concentrate band removed at the ports is controlled by adjustable splitters. The

grade of concentrate taken from descending ports progressively decreases,

tailings being discharged from the lower end of the spiral conduit. Spirals are

made with shapes of varying steepness, the angle affecting the specific gravity of

separation, but having little effect on the concentrate grade and recovery.

Figure 2.9 Cross-section of spiral stream

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2.5.1.2 Shaking Tables

The principle behind the shaking table is that when a flowing film of water flows

over a flat, inclined surface the water closest to the surface is retarded by the

friction of the water absorbed on the surface; the velocity increases towards the

water surface [5]. If mineral particles are introduced into the film, small particles

will not move as rapidly as large particles, since they will be submerged in

slower-moving portion of the film. Particles of high specific gravity move more

slowly than lighter particles, and so a lateral displacement of the material will be

produced (fig. 2.10). The flowing film effectively separates coarse light particles

from small dense particles, and this mechanism is utilised to some extent in the

shaking-table concentrator (fig. 2.11).

Figure 2.10 Action in a flowing film

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Figure 2.11 Shaking table

The shaking-table concentrator consists of a slightly inclined deck, A, onto which

feed at about 25% solids by weight, is introduced at the feed box and is

distributed along C; wash water is distributed along the balance of the feed side

from launder D. The table is vibrated longitudinally by the mechanism B, using a

slow forward stroke and a rapid return, which causes the mineral particles to

crawl along the deck parallel to the direction of the motion. The minerals are

thus subjected to two forces-that due to the table motion and that at right angles

to it due to the flowing film of water. The net effect is that the particles move

diagonally across the deck from the feed end and, since the effect of the flowing

film depends on the size and density of the particles, they will fan out on the

table, the smaller, denser particles riding highest towards the concentrate

launder, which runs along the length of the table. Fig. 2.12 shows an idealized

diagram of the distribution of table products. An adjustable splitter at the

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concentrate end is often used to separate this product into two fractions: a high-

grade concentrate and a middlings fraction.

Particle size plays a very important role in table separations; as the range of

sizes in a table feed increases, the efficiency of separation decreases. If a table

feed is made up of a wide range of particle sizes, some of these sizes will be

cleaned inefficiently. Thus since the shaking table effectively separates coarse

light from fine dense particles, it is common practice to classify the feed since

classifiers put such particles into the same product, on the basis of their equal

settling rates. This classification can be effected through use of classifiers such

as multi-spigot hydrosizers, and such other classified as described in section 2.4.

The capacity of a table varies according to size of feed particles and the

concentration criteria. Tables can handle up to 2 tonnes per hour of 1.5mm sand

and perhaps 1t/h of fine sand. On 100- PIHHGPDWHULDOVWDEOHFDSDFLWLHV

may be as low as 0.5t/h.

The quantity of water used in the feed pulp varies but for ore tables normal feed

dilution is 20-25% solids by weight. In addition to the water in the feed pulp, clear

water flows over the table for final concentrate cleaning. This varies from a few

litres to almost 100 litres/min according to the nature of the feed material.

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Tables slope from the feed to the tailings discharge end and the correct angle of

incline is obtained by means of a hand wheel. The table is slightly elevated along

the line of motion from the feed end to the concentrate end. The correct amount

of end elevation varies with feed size and is greatest for the coarsest and highest

gravity feeds. Normal end elevations in ore tabling range from a maximum of

90mm for a very heavy coarse sand to as little as 6mm for an extremely fine

feed.

2.5.1.3 The Mozley Laboratory Separator

This flowing film device (fig. 2.14), which uses orbital shear, is now used in heavy

mineral processing laboratories, and is designed to treat small, samples (100g of

ore) [11]. The separator consists of a v-shaped stainless steel tray measuring

128cm in length, 72cm in width, 91cm in height and weighing 150kg. Below the

tray is the shaking mechanism, which causes the tray to vibrate longitudinally at

amplitude of inch at a frequency of 120 to 240 rpm. The transverse cyclic

oscillation is at 60 to 120 rpm. At one end of the tray is the feed cone and

concentrate wash water pipe, while an irrigation water pipe runs along the

length of the tray with holes perforated on the pipe at equal separations, from

which jets of water impinge on the sample. At the other end of the tray is the

tailings launder for collection of tails. The tray is sloped at 1-50 to the horizontal

from the feed end. This laboratory piece of equipment has the advantage that its

V profile tray with end knock, when treating closely sized material, is capable

not only of duplicating heavy liquid analysis results, but of giving additional data

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in the higher specific gravity ranges. It is also able to predict shaking table

performance when treating hydraulically classified products. The separator is

therefore useful in carrying out release analysis and for prediction of reaction of

feed to shaking table concentrator, and the efficiencies of such processes in

gravity separation [5].

Figure 2.14 Schematic diagram of a Mozley laboratory separator

2. 5. 1. 4 Cent rif ugal Concent rators

A number of centrifugal gravity separation devices, designed to treat ultra-fine

particles, are available for the gravity separation of a mixture of ground alloy and

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D. Muchena 2003 28

gangue in slag. The Mozley Multi-Gravity Separator (MGS) shall be considered in

this section.

The principle of the MGS can be visualized as rolling the horizontal surface of a

conventional shaking table into a drum, then rotating it so that many times the

normal gravitational pull can be exerted on the mineral particles as they flow in

the water layer across the surface [12]. Fig. 2.15 shows a cross-section of the

pilot scale MGS.

Figure 2.15 Pilot scale MGS (Courtesy Natgroup)

The plant-scale MGS consists of two slightly tapered open-handed drums,

mounted back to back, rotating at speeds variable between 90 and 150rpm,

enabling forces of between 5 and 15g to be generated at the drum surfaces. A

sinusoidal shake with an amplitude variable between 4 and 6cps is

superimposed on the motion of the drum; the shake imparted to one drum being

balanced by the shake imparted to the other, thus balancing the whole machine.

A scraper assembly is mounted within each drum on a separate concentric shaft,

driven slightly faster than the drum but in the same direction. This scrapes the

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D. Muchena 2003 29

settled solids up the slope of the drum, during which time they are subjected to

counter-current washing before being discharged as concentrate at the open,

outer, narrow end of the drum. The lower density minerals, along with the

majority of the wash water, flow downstream to discharge as tailings via slots at

the inner end of each drum.

2. 6 Assessment of Met allurgical Ef f iciency

In the concentration of alloy in the Sirosmelt slag, the recoveryis defined as the

percentage of the total metal contained in the slag that is recovered in the

concentrate. Thus a recovery of 90% will mean that 90% of the alloy in the slag is

recovered in the concentrate and 10% lost to the tailings. Recovery is usually

expressed in terms of the valuable end product. The ratio of concentrationis the

ratio of the weight of the feed to the weight of the concentrates. It is a measure of

the efficiency of the concentration process and it is closely related to the gradeor

assayof the concentrate. The value of the ratio of concentration will generally

increase with the grade of concentrate. The grade, or assay, usually refers to the

content of the marketable end product in the material.

The enrichment ratiois the ratio of the grade of concentrate to the grade of the

feed and again is related to the efficiency of the process. Ratio of concentration

and recovery are essentially independent of each other, and in order to evaluate

a given operation it is necessary to know both. There is an approximately inverse

relationship between recovery and grade of concentrate in all concentrating

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D. Muchena 2003 30

processes. If an attempt is made to attain a very high-grade concentrate, the

tailings assays will be higher and the recovery low. If a high recovery of metal is

aimed for, there will be more gangue in the concentrate and grade of concentrate

and ratio of concentration will both decrease. It is impossible to give figures for

representative values of recoveries and ratios of concentration. The aim of milling

operations is to maintain the values of ratio of concentration and recovery as high

as possible, all factors being considered.

Since concentrate grade and recovery are metallurgical factors, the metallurgical

efficiencyof any concentration operation can be expressed by a curve showing

the recovery attainable for any value of concentrate grade. A typical recovery-

gradecurveshowing the characteristic inverse relationship between recovery

and concentrate grade is shown in fig. 2.16.

Fig. 2.16 Recovery and concentrate grade

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Concentrate grade and recovery, used simultaneously, are the most widely

accepted measures of assessing metallurgical performance; thus the two

measures will be used in this project to assess metallurgical performance of

gravity separation processes to recover alloy from slag. These two measures can

be combined into a single index defining metallurgical efficiency of the

separation:

gm RRES =..

Where S.E.=Separation efficiency

Rm = % recovery of the valuable mineral

Rg = % recovery of the gangue into the concentrate

Ff

Cc

Rm%100

=

The gangue content recovery of the concentrate = ( )%100100 mc

Where m is the percentage metal content of the valuable mineral, i.e.

Gangue content = ( ) mcm 100 ,

Rg = C x gangue content of concentrate/gangue content of feed

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Rg =( ) fmcmC 100

Therefore, ( ) ( ){ }fmcmCf

CcRR gm

=

100

100

( )

( )ffm

fcCmRR gm

=

100

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3. 0 EXPERI MENTAL I NVESTI GATI ONS

3. 1 Methodology

Firstly, a series of mineralogical analysis of the slag, such as chemical

analysis, sieve tests and metallographic analysis, will have to be conducted.

Mineralogical analysis of the slag, which involves characterization of samples

of the slag, was carried out, aimed at establishing the following:

a) Slag to metal/alloy proportions,

b) Mode of occurrence of entrapped copper and nickel,

c) Compositional variations between samples.

It was hoped that this exercise would also offer an opportunity to investigate

the degree of grinding required for effective liberation of the values from the

gangue, and effectiveness of the separation methods in concentrating copper

and nickel [13]. Information from the mineralogical analysis will then be used

to design a preliminary experimental flow sheet for laboratory scale, milling

and physical separation investigations to predict the behaviour of the slag to

such techniques. Laboratory scale separation will be carried out. Results from

these tests will lead to design of a plant/pilot plant scale flow diagram for the

milling of, and separation of the alloy from the slag.

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3. 2 Chemical Analysis

Chemical Analysis was done to establish the elements in the slag and their

relative abundances. This was accomplished through collection of statistics from

E.N.R. Sirosmelt plant assay record sheets for the period from 01/09/2002 to

25/09/2002 from the daily product accounting samples and the average assays

for different components from the slag-tap were calculated. The results are given

in 4.1.

3.3 Metallographic Analysis

Two polished sections were prepared from slag samples by being first ground to

flat surfaces, then being successively ground and polished until a mirror finish

was obtained for each surface. The sections were studied using a Zeiss reflected

light microscope and photographs were taken. Results obtained which are meant

to establish the mineralogy of the slag, are shown in 4.2.

3. 4 Part icle Size Analysis

A 4.5kg sample of the slag was crushed in a laboratory jaw crusher then ground

to passing 1mm. A sieve analysis was carried out in the size range plus 1000m

to minus 53m. The sieves used were the following: 1000m, 850m, 500m,

425m, 355m, 212m, 125m, 75m, 53m aperture sieves. Each of the 9 size

fractions was assayed for copper and nickel content and the metal distribution

was computed. The results of the metal distribution are given in section 4.3.

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3. 5 Recovery versus Part icle Size Analysis

A Mozley Multi-gravity Laboratory Separator was chosen as the technology for

separation. The operation of this separator is explained in 2.4.1.3. 100g of slag

were crushed for different size fractions in the range minus 1000 to plus

53microns and collected using sieves. The sieves used were the following:

1000m, 850m, 500m, 425m, 355m, 212m, 125m, 75m, 53m aperture

sieves. The 100g sample of slag crushed to a particular size was placed on a

Mozley Multi-gravity Laboratory Separator and wetted. The cyclic motion of the

tray mobilised the slag particles enabling stratification to take place. The heavy

metal particles sank to the tray surface and were thrown upstream by the end-

knock action. The lighter (gangue) mineral particles were carried downstream by

the flow of irrigation water to discharge via the tailings launder.

Interpretation of the data was carried out by assay analysis of the separator

products. For each particle size fraction, concentrates were collected and sent for

assaying for Cu and Ni. The assay results were used to calculate the recovery

(see appendix 3) for each particle size fraction and recovery versus particle size

curves drawn (see 4.4).

3. 6 Gravit y Concent rat ion Test work on Sirosmelt Slag

Samples were collected from different parts of the slag dump in varying amounts.

The samples were crushed in a laboratory jaw crusher and ground in a cone

crusher to pass the 850m screen. The grinding product was mixed

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homogeneously by coning and quartering to obtain a 3kg sample. A Mozley

Multi-gravity Separator Table was used for gravity separation (see 2.4.1.3). A

flow sheet was first designed for the gravity separation program to be followed.

This was done in order to achieve the maximum recoveries possible taking into

account that a satisfactory recovery cannot be possible to achieve in one pass.

Thus the sample would have to be subjected to a series of gravity separation

process runs through the Mozley Gravity Separator Table. Thus a rougher-

scavenger-cleaner-recleaner type of flow sheet (fig. 3.1) was designed, taking

into account its efficiency in achieving high recoveries without compromising the

grade of the concentrate.

A 3kg sample of crushed slag was fed dry into the feed cone at one end of the

Mozley Multi-gravity Separator, (MMGS) table under a running stream of water to

form a pulp of about 25% solids. The MMGS table was turned on and

simultaneously the water through the irrigation and concentrate pipes was

turned on. The feed cone was unscrewed to let the feed flow as a pulp onto the

table. The pulp was sprayed with a stream of water from the irrigation and

concentrate wash pipes to enhance flow. The table was sloping towards the

tailings end at approximately 5o. The sample was subjected to the gravity

separation action for one minute after which the machine was switched off. The

slag ground to passing 850m and was fed to the MMGS table. The concentrate

from the rougher separation run, MZT1, was collected, dried and weighed and

was subjected to another run which is the cleaner separation run, MZT2 (after a

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D. Muchena 2003 37

small measured amount was taken for assaying purposes). The concentrate from

MZT2 was taken for the recleaner separation run, MZT4 to give separate products

of concentrate and tailings. The MZT2 tailings were collected, to be re-fed to

MZT1 together with fresh feed, while the concentrates were treated as the final

concentrate, dried and weighed and taken for assaying. The tailings from MZT1

was further ground to 80% passing 125m and subjected to a scavenger

separation run on the MMGS, MZT3 under the same conditions as described

above. The tailings from MZT3 were collected as the final tailings, dried and

weighed and taken for assaying while the concentrate was collected, to be re-fed

to MZT1 together with fresh feed. Two products were obtained for each run with

the lighter gangue minerals being collected at the tailings end while the heavier

value metal was obtained at the concentrate end. The products were decanted

and oven dried at 105oC for 2hours. A sample was extracted from each product

and taken for assaying. The results for the above test work are shown in 4.5.

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Fig 3.1: Schematic experimental flow sheet for the concentration of alloy from slag

Slag Feed

850m

MZT1

MZT2Grind to 80% passing 125 m

MZT3MZT4

concentrate tailings

concentrate

Final tailing

tailings

Finalconcentrate

concentrate

tailings

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3. 7 Spiral Concent rat ion Test work on Sirosmelt Slag

A 20kg sample of crushed Sirosmelt slag was mixed with water to form a pulp of

density 40% solids and was pumped into a spiral concentrator, at the top. As the

pulp flowed spirally downwards stratification of particles due to centrifugal force

occurred. Two separate products were obtained at the concentrates and tailings

ports. The products were dried and weighed and taken to assay for copper and

nickel. The results obtained are summarized in 4.6.

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4. 0 RESULTS AND DI SCUSSI ON

4.1 Chemical Analysis

TABLE 4.1:CHEMICAL ANALYSIS OF SAMPLES OF THE SLAG

%Ni %Cu %S %Fe %FeO %Co %CaO %SiO2 %Al2O3 %MgO %Cr2O30.97 1.12 0.10 10.06 12.06 0.36 28.80 26.70 11.74 5.70 1.84

The results above show that the average assays for Cu and Ni are above

the expected limits of a total of 1% for both, at 0.97% and 1.12%

respectively, totaling 2.09%. This high metal-slag loss situation can be

remedied through a mineral processing technique.

4. 2 Metallographic Analysis

The results of metallographic analysis of slag samples are shown in figs. 4.1a

and 4.1b.

Figs 4.1b Fig. 4.1b

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Slag samples showing different phases present in the slag. [Magnification: 50x]

After careful study and analysis of the phases shown above, the following

deductions were made:

i. Most of the copper and Nickel in the slag occurs as an alloy within matte

granules. Minor hosting phases of copper and nickel include chalcocite

(Cu2S), cuprite (Cu2O), tenorite (CuO), bunsenite (NiO) and covellite (CuS),

in order of decreasing abundance [14].

ii. Generally, Au, Pd, and Pt correlate with Cu and Ni. This indicates that the

three are associated with Cu and Ni in the alloy. This could be confirmed by

microbe analysis.

iii. Slag fragments make up the ultimate majority of the samples. The slag

particles are mainly massive silicate intercepted by carbonate and iron

oxide, mainly magnetite. Fine matte particles (< 75m) occur as rounded

globules in relatively coarse fragments of slag.

iv. Majorities of the entrained matte particles have clear margins and only rare

cases of coating were observed. The coatings are broken rims of cuprite

and/or magnetite. The rest of the magnetite occurs mainly as spherical and

rod-shaped globules and less commonly as euhedral particles in the slag.

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D. Muchena 2003 42

v. Free matte granules occur as rounded to subangular particles, most of them

free of oxidation rims. Majority range in size from 600 to 1000m, average

size 300m. Patches of cuprite were visible inside slag. Finer particles of

cuprite,

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From the results shown, it was observed that metal distribution in the various size

ranges shows that 92.77% of the copper and 91.09% of the nickel is contained in

the size fraction minus 1000 to plus 125microns. This can be seen from the

graph in fig. 4.2 where the size range is the steepest part of the graph. The

effectiveness of any separation process outside this range becomes less.

Particles greater than 1000micronsconstitute 49.62% of the copper and 37.45%

of the nickel respectively, while particles less than 125micronsconstitute 7.23%

of the copper and 8.91% of the nickel. Therefore to effectively liberate the

particles and still save on comminution costs the particles should be ground in

the above size range.

4. 4 Recovery versus Part icle Size Analysis

Fig 4.3: Recovery Vs Particle Size for Copper

0

10

20

30

40

50

60

70

80

90

1000 850 500 355 212 125 75 53 53-

size (microns)

Recovery(

%),Grade(%)

grade

recovery

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D. Muchena 2003 44

The performance of the separator varied with the feed size as shown on the

graph in fig. 4.3 and fig 4.4. Recoveries of copper and nickel by size show that

plus 60%recoveries occur between minus 850micronsto plus 125microns. The

effectiveness of the gravity separation process falls for particles below

125micronsdue to loss of values in fines at fine sizes. The fall at 500micronsis

probably due to loss of values with gangue due to little unlocking at that size for

metal particles entrained in the slag at sizes lower than 850microns. This shows

that in a circuit to be designed for the recovery of metal values from E.N.R. slag,

the first gravity separation run should be performed on particles ground to

Fig 4.4: Recovery vs Particle Size for Nickel

0

10

20

30

40

50

60

70

80

90

1000 850 500 355 212 125 75 53 53-

size (microns)

recovery(%),grade(%)

grade

recovery

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D. Muchena 2003 45

passing 850microns, whereupon the tailings from this run would be further

ground to passing 125micronsthen subjected to another gravity separation run.

4. 5 Gravit y Concent rat ion Test work on Sirosmelt Slag

The results are summarised in the experimental flow sheet in fig. 4.5

Generally high recoveries in the concentrate were obtained for each of the

stages above. The final separation stage MZT4 gave the highest recoveries of

Nickel and Copper in the concentrate, of 88.71% and 92.34% respectively. At

this stage it became unnecessary to continue with further separation as high

enough recoveries were obtained. A high grade of the alloy, totalling 85.62% was

also obtained. Thus it can be safely concluded that further concentration beyond

this stage in not necessary.

The concentrate of the alloy obtained from gravity separation can be recharged

together with pellets into the Sirosmelt furnace or added to the granulated matte,

and fed to the siroleach plant.

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Fig 4.5: Experimental flow sheet

Feed

wt %

% Cu, %Ni

Rec (Ni, Cu)

MZT1

MZT2

MZT4

MZT3

wt %

% Cu, %Ni

% Cu, %Ni

% Cu, %Ni% Cu, %Ni

% Cu, %Ni

% Cu, %Ni

% Cu, %Ni

wt %

wt %

wt %

wt %

wt %

wt %

Rec (Ni, Cu)

Rec (Ni, Cu)

Rec (Ni, Cu)

Rec (Ni, Cu)

Rec (Ni, Cu)

Rec (Ni, Cu)

Rec (Ni, Cu)

wt %

% Cu, %Ni

Rec (Ni, Cu)

100.00

1.12, 0.97

100.00, 100.00

concentrate tailing

Grind

80%,

125m

tailing concentrate

Final concentrate tailing

Final tailing concentrate

70.36

13.27, 2.35

48.15, 46.67

29.63

36.00, 6.02

51.85, 53.33

49.64

22.41, 5.81

47.88, 30.90

50.36

49.40, 6.23

52.12, 69.11

35.93

51.40, 34.22

88.71, 92.34

64.07

1.20, 0.61

11.29, 7.66

6.11

37.40, 32.86

55.01, 57.13

93.89

1.83, 1.75

44.99, 42.87

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4. 6 Spiral Concent rat ion Test work on Sirosmelt Slag

Table 4.4: Spiral Concentration of Sirosmelt Slag

The concentration grade is low, at 6.74% for Cu and 6.13% for Ni. However the

recoveries are high at 92.86% and 97.51% for copper and Nickel respectively.

Thus although little metal is lost to tailings, the resultant grade after spiral

concentration is low.

Sample wt wt Assay Cu Assay Ni Recovery Recoveryg % % % Cu % Ni %

Spiral Feed 20000 100 1.12 0.97 100.00 100

Spiral Conc 3086 15.43 6.74 6.13 92.86 97.51124

Spiral Tail 16914 84.57 0.0945 0.028546 7.14 2.48876

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5. 0 CONCLUSI ONS

1. The average assays for Cu and Ni are above the expected limits of a total

of 1% for both, at 0.97% and 1.12% respectively, totalling 2.09%. This

high metal-slag loss situation can be remedied through a mineral

processing technique.

2. Since metallic species in the slag occur in the size range from 1000m, grinding at 1mm can be envisaged to liberate majority of the

matte particles coarser than 200m. For particles less than 15m, very

fine grinding can unlock them.

3. The large difference in specific gravities between the Ni-Cu alloy and the

slag, which gives a correspondingly high concentration criterion, can be

manipulated in choosing a physical separation technique for concentrating

the Cu-Ni alloy. The appropriate technique chosen was gravity separation,

which utilises differences in specific gravities between materials in

separating them.

4. Metal distribution in the various size ranges shows that 92.77% of the

copper and 91.09% of the nickel is contained in the size fraction minus

1000 to plus 125microns. Therefore to effectively liberate the particles and

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D. Muchena 2003 49

still save on comminution costs the particles should be ground in the

above size range.

5. Recoveries of copper and nickel by size show that plus 60%recoveries

occur between minus 850micronsto plus 125microns. The effectiveness

of the gravity separation process falls for particles below 125micronsdue

to loss of values in fines at fine sizes.

6. The Mozley multi gravity separator table produced high recoveries for both

copper and nickel in the final concentrate, at 92.34% and 88.71%

respectively, after subjecting the slag to a rougher-scavenger-cleaner-

recleaner gravity concentration circuit. The grades for copper and nickel

were 51.40% and 34.22 for the same process. Thus it can be concluded

that shaking tables can be used in the concentration of the alloy in the

slag.

7. The spiral concentrator produced very low grades of the alloy although

recovery was high. This shows that little separation between the gangue

and the alloy was effected. Thus their use in concentrating the alloy was

less effective.

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6. 0 RECOMMENDATI ONS

6.1 In a plant production scenario milling can be effected through

crushing the slag using a jaw crusher and using a Barmac crusher

for size reduction down to passing 1mm. A complete milling circuit

would encompass all the aspects like screening and sizing.

6.2 The laboratory scale gravity separation equipment used is useful in

duplicating gravity separation behaviour of equipment like James

shaking table. But due to the gradual phasing out of the James

shaking table as a gravity separation device over the past years,

such equipment as spirals and the Mozley C902 Multi Gravity

Separator Drum are highly recommended for alloy from slag (AFS)

applications.

6.3 The product from gravity separation is of a high grade at high

recoveries. Thus further concentration beyond gravity separation is

not recommended considering costs of operation versus

improvement in grade.

6.4 A spiral concentrator flow sheet incorporating many spirals in a

separating circuit needs to be designed in order to maximise the

upgrading exercise, for a higher grade of alloy to be obtained from

spiral concentration.

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6.5 The product from gravity separation can be fed into the Sirosmelt

furnace together with pellets or fed to the siroleach together with

granulated matte.

6.6 The recommended plant flow diagram for the whole process of

recovering Ni-Cu alloy from slag is shown in appendix 6.

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7. 0 BI BLI OGRAPHY

1. RioZim E.N.R., Empress Nickel Refinery Operations Manual, RioZim

E.N.R., Zimbabwe, 2001, pp1-5

2. Hausen, D.M., Evaluation and Optimisation of Metallurgical

Performance, SME Inc, 1991, Chapter 17

3. Bernhardt, C., Particle Size Analysis, Chapman and Hall, London, 1994,

pp8-26

4. Anon. Test Sieving, British Standard 1796, London Press, London, 1976,

pp9-11

5. Wills B.A., Mineral Processing Technology, Sixth Edition, Pergamon

Press, Oxford, 1996, pp7-8, 116-124, 142-176

6. Lewis, F.M. et al. Comminution: A guide to size reduction system

design. Min. Eng., 1976, pp28, 54-55

7. Rodriguez, D.E. The Tidco Barmac Autogenous Crushing Mill-a circuit

design primer, Minerals Engineering, 1990, pp53-54

8. Taggart, A.F. Handbook of Mineral Dressing, Wiley, New York, (1945),

pp213-215

9. Pearse, G. Some Manufacturers of hydrocyclones, Mining Magazine,

1988, pp106

10. Mills, C., Mineral Processing Plant Design, AIMME, 1978, pp38-39

11. Wills, B.A., Laboratory simulation of shaking table performance,

Mineral magazine, 1981, pp87-89

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D. Muchena 2003 53

12. Mozley R. et al., The Mozley Mineral Separating Systems, 2000, pp2

13. Evans, A.M., An Introduction to Ore Geology, Blackwell Scientific

Publications, Oxford, 1980, pp5-6

14. Craig, J.R., Vaughan, J.D., Ore Microscopy and Petrography, John

Wiley and Sons, Canada, 1981, pp1-10

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8. 0 APPENDI CES

Appendix 1: Calculation Of Total Metal-Slag Losses

Total tonnage of slag to date= 11958.3 tons

Total amount of copper lost to slag= 0.97% of 11958.3 tons = 115.99551tons

Total amount on Nickel lost to slag= 1.12% of 11958.3 tons = 133.93296 tons

Total amount on alloy lost to slag= 250 tons

Total revenue lost due to copper lost to slag= U.S.$1,600,00 x 116 =U.S.$185,592,82

Total revenue lost due to nickel lost to slag=U.S.$7,400,00 x 134 =U.S.$991,103,90

Total revenue lost due to Ni + Cu lost to slag= U.S.$1,176,696,72

Appendix 2: Calculation of Concentration Criterion

).(.).(.

).(.).(.

mFluidMediuGSeralLighterMinGS

eralLighterMinGSeralHeavierMinGSonionCriteriConcentrat

=

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Appendix 3: Calculation of Recovery

Ff

Ccery

%100covRe

= or

Ff

Ttery

%100covRe

=

WhereC = wt % of concentrate,c = grade of concentrate,T = wt % of tailingt = grade of tailingF = wt % of Feed,F = grade of feed

Appendix 4: Copper And Nickel Distribution In The SizeFractions

Size Mass Wt Assay

Assay

Distribution

Distribution

Cumulative Cumulative

(Microns)

g (%) Cu(%)

Ni(%)

Cu (%) Ni (%) Distribution Cu(%)

DistributionNi (%)

1000 1219.5

27.05 2.11 3.44 49.62 37.45 49.62 37.45

850 292.3 6.48 0.67 1.91 3.78 4.98 53.39 42.43500 669.9 14.86 1.37 2.61 17.72 15.61 71.11 58.04425 381.1 8.45 0.97 2.25 7.16 7.65 78.27 65.69355 243.6 5.40 0.76 2.49 3.55 5.41 81.82 71.11212 657.5 14.59 0.69 2.69 8.75 15.79 90.57 86.90125 299.3 6.64 0.38 1.57 2.19 4.19 92.77 91.0975 354.1 7.86 0.50 1.13 3.40 3.57 96.16 94.6653 210 4.66 0.39 1.45 1.60 2.72 97.76 97.38-53 180.3 4.00 0.65 1.63 2.24 2.62 100.00 100.00

Total 4507.6

100.00 1.12 0.97 100.00 100.00 100.00 100

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Appendix 5: Effect of Particle Size on Gravity Separation

Size Concentrate

(Microns) mass wt Assay Assay Recovery

Recovery

Recovery

g % Cu % Ni % Cu % Ni % (Cu+Ni)%

1000 8.8 4.4 20.4 14.44 5.317536

18.46977

7.544094

850 17.4 8.7 39 13.32 63.42056

60.67225

62.69752

500 19.2 9.6 37.4 14.02 33 51.56782

36.59244

355 19.7 9.85 30.8 14.86 50.14545

58.78353

52.66405

212 27.4 13.7 25.2 15.14 63.81516

77.10706

68.22938

125 14.3 7.15 35.8 14.72 84.47855

67.03694

78.52565

75 8.5 4.25 34.4 14.72 36.73367

55.36283

40.85323

53 2.3 1.15 43.4 14.58 15.84444

11.56345

14.495

53- 1.7 0.85 42.2 7.15 6.97859

9

3.72852

8

6.19608

6Total 119.3 59.65

8/8/2019 Pages 1 57

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Appendix 6: Suggested Alloy From Slag (AFS) Plant flow Diagram

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