In ternatio nal Journal o f Mineral Processing, 10( 1983) 183--204 183 Elsevier Scientific Publishing Company, Amsterdam -- Printed in The Netherlands

BENEFICIATION OF IRON ORE BY FLOTATION - - REVIEW OF INDUSTRIAL AND POTENTIAL APPLICATIONS

R. HOUOT

Centre de Recherches sur la Valorisat ion des Min$rais de I 'E.N.S.G., L .A . 235, B.P. 40 -- 54501 Vandoeuvre- les -Nancy CSdex (France)

(Received April 23, 1982; revised and accepted September 30, 1982)

ABSTRACT

Houot, R., 1983. Beneficiation of iron ore by flotation -- review of industrial and poten- tial applications. Int. J. Miner. Process., 10: 183--204.

The market requirements for higher-grade concentrates of iron to improve the produc- tivity of the iron and steel industry, has increased the importance of the flotation process with respect to the conventional preconcentration of ore by gravity or magnetic separa- tion. The flotation method most commonly applied is the one that is based on cationic flotation of silica and silicates (reverse flotation), and which is preceded, or not, by desliming or selective flocculation.

INTRODUCTION

Research on the flotation of iron ore, on either a bench or a pilot-plant scale, started towards 1931 (Crabtree and Vincent, 1962). It was mostly confined to ores containing siliceous gangue, which appears to be the most abundant type in the world and the reserves of which, when based on a minimum tenor of 30 to 35% iron, amounts to a respectable quanti ty of milliards of tons.

The first integration of flotation in a plant dates from 1954 (Humboldt Mine). The second one started in 1956 (Republic Mines).

The four plants in the United States, which adopted flotation (Merril and Pennington, 1962) totalled only 5700 t /d in 1960. In 1974 (Anonymous), that country showed a crude ore capacity of approximately 130,000 t/d, the ore of which was entirely or partly subjected to concentration by flotation.

At the present {1980), about 120 million t /y of crude ore is treated by this technique in the western world.

Primarily, the development of the technology was (up to 1960) essentially of North American origin [studies from companies: Hanna Mining Company (Bunge and Pocrnich, 1960); Cleveland Cliffs Iron Company (Leach et al., 1977) ; U.S. Bureau of Mines] but was then modified in a great number of

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184

countries in which specific mining problems led to the study of the subject by numerous teams.

TREATED ORES (Routhier, 1963)

The ores that are treated are essentially of the jaspilite and ferruginous quartzite type. They are known by various names: magnetite quartzite or/and specular hematite; ferriferous jaspilite; "banded iron stones"; taconite; and itabirite. They constitute the largest iron reserves of the world and at the present, those areas which were enriched by natural processes are actively exploited. It is chiefly the lower-grade ore that is submitted to stages of preconcentration (gravity, magnetism, flotation). The sediments were deposited in ancient marine environments {usually of Precambrian age) and all were more or less subjected to metamorphism.

The minerals encountered are: {a) magnetite-hematite or specular hematite - -mar t i t e {magnetite substi tuted epigenetically by hematite); (b) quartz mosaic; and (c) various silicates: amphiboles of variable iron content, epidote -- garnet -- diopside -- micas.

However, according to recent studies, the degree of oxidation of the iron oxides (hematite, magnetite) is not related to the metamorphic grade but to the paleogeography of the basins*.

The enrichment of certain portions of the deposits is essentially due to supergene alteration by meteoric solutions, which promoted the leaching of silica. In the Brazilian ores, this alteration is observed over several tens of metres: from itabirites containing 25 to 40% iron, zones formed which were enriched to 66% iron because of the substitution of silica by hematite.

PROCESSES USED

Historically, methods of direct flotation of iron oxides by anionic reagents were the first to be tested and applied. The following reagents were used:

- - sulphonates of petroleum + fuel - - fat ty acids (Hanna float and derived processes) - - Dual process -- hydroxamates

These techniques were then followed by the application of reverse flotation of silica and silicates, the ferriferous concentration being collected at the bo t tom of the cell.

*The model of H.L. James (1954) could be applied to this type of basin, which would then explain the virtual absence of carbonates. According to the general diagram of Krumbein and Garrels (Garrels and Christ, 1965) the depositional environment occurred close to the surface and was sufficiently oxygenated to destroy the organic matter (pyrite is therefore absent) and for the iron to be present in the form of magnetite (at moderate depth) or hematite (at shallow depth in the basin).

185

The following routes were tested successively: (a) anionic flotation of the activated silica; {b) cationic flotation of the silica: and (c} selective floccula- tion of the ferriferous slimes with respect to the silicate slimes which are eliminated by overflow, associated with the anionic or cationic flotation of silica. Most of these processes took shape in one or more flotation plants, some of which were since closed down because the deposit was worked out, while others were converted to different methods because of change in the composit ion or in the grade of the ore and the grain size required for the concentrates.

In general, the market requirements for the iron grade of the concentrates tend to increase and at the present they are approximately 66% Fe for utili- zation in the traditional iron and steel industry (blast furnace) after pelletiza- tion, whereas the SiO2 content should not exceed 2% (68 to 70% Fe according to the mineralogy) if these concentrates are to be used in the direct prereduction (or reduction) route-electric steel works.

F L O T A T I O N WITH P E T R O L E U M S U L P H O N A T E

The first results obtained by this type of reagent led to a series of patents taken out by E.C. Herkenhoff {1945} for the account of the Cyanamid Com- pany.

The flotation is done at acid conditions (pH -~ 3) in the presence of fuel and sulphonate. Desliming is indispensable to obtain good results. The con- ditioning is done in a thick pulp (65 to 70% solids).

Investigation of the petroleum sulphonate flotation procedure began in 1949 in the Hanna Mining Company. It was found that a small addition of fat ty acid improved the metallurgical performance. The ore selected, ori- ginating from the deposit of Groveland (Michigan), was first upgraded by gravimetry (3 stages of spirals}. The gravimetric tailings const i tuted the feed for the flotation. The evolution of that plant {15,000 t /d) is typical for the imposed conditions {Smith and Sougstad, 1962; Maumen~, 1965; Bunge et al., 1977).

Phase 1 : Gravity flowsheet {spirals) -- flotation of gravimetric tailings. Table I and Fig. 2 illustrate the results obtained.

Phase 2: Flowsheet of integral flotation {line 4). With the same reagents, the integral flotation permitted of obtaining weight efficiencies of 42,3% of a concentrate assaying 64% iron with a recovery close to 77%.

Phase 3: With the increase in magnetism at greater depth of mining, it became apparent that the introduction of magnetic separation would be attractive. For that reason a new flowsheet was adopted, which is given in Fig. 3; the results are summarized in Table I.

In none of these instances did flotation of silica and silicates prove to be of economic value because of the mineralogy. The amorphous forms of silica (chert, jasper) and the various silicates (garnets and amphiboles} indeed limited the flotation to levels inferior to those obtained with anionic flota- tion.

186

p H ~

÷0.I \ ~

",~,;

3 0.0' oR c

PEAT Eh

-0.3

PEAT PYRI1[

HEMATITE LIMONITE MN OXIDES SILIC& CHAMOSITE" Calcite Phosphorite

ZO 8.0 I

CALCITE [S-AUN'FY'200°I--I H~o~I~ IGYPSUM I ~i;~ s IANHYOR/TEI C~,,osite IHALITE I Phosphorite IDol OMI TE I Silica lE rc " j

SILICA Rhodochrosite Alabondite (?) Calcite Primary uranium concent rat ions Primary heavy metal

Sulphldes

GAN_tC . MATTER . FENCE (Eh. O)

Hematite /:1 ,mooit, I-.4NHYORITE Cha mosite IDOLOMITE

CHAMOSITE \ '~ontelo~,cA.,cM~rrD SIDERITE " , ~ L erc GL&UGONITE C ~ RIQOOOGHROSITE ORGANIC

MATTER SILICA PHOSPHORITE

Ca,c,te \ \ Primary uranium

Siderite r LIN 200°/o4 RhodochrositelGY,~$UM Phosphorite IANHYORITE |

~-~ ~ ~GAN/¢ MATTE

CALCITE ORGANIC MATTER

Pyrite [SALiNlrY > 200 Olooi Phosphorite [Gypsum ,"_'~, I Aloband[te (?)lAnhydr~ ~ I I

l p r m r t : " J - ,

'Charnos~te as used here is representative at the sed;mentary iron silicates

Fig. 1. Sedimentary chemical end-member associations in their relations to environmental limitations imposed by selected Eh and pH values.

This type of f lotation has not brought forth any other new achievements. Research is virtually absent in this direction (Hancock, 1974) .

BENEFICIATION WITH FATTY ACIDS

Flotation with sulphonates soon had to make way for f lotation with fatty acids. Thus, the washing plant of Humboldt (Maumen~, 1965; Leach et al.,

T A B L E I

Mater ia l ba lance in Groveland. Top: Spi ra l - f lo ta t ion circui t (1961 ). B o t t o m : f l o t a t i on circui t ( 1 9 7 2 )

187

Magnetic-

P r o d u c t s % weigh t % Fe % SiO2 % Fe d i s t r ibu t ion

Spiral c o n c e n t r a t e 32.7 60.7 10.1 59.6 F l o t a t i o n c o n c e n t r a t e 13.9 58.2 8.9 24.3 Tail ings 53.4 10.0 16.1 Tota l c o n c e n t r a t e 46.6 60.0 9.8 83.9

Feed 100.0 33.3 100.0

P roduc t s % weigh t % Fe % SiO2 % Fe d i s t r ibu t ion

Magnet ic c o n c e n t r a t e 27.1 F l o t a t i o n c o n c e n t r a t e 21.8 Tota l c o n c e n t r a t e 48.9 Feed E lu t r i a te c o n c e n t r a t e 42.1

Reagents of first c o n d i t i o n e r : s u l p h o n a t e fuel no. 2 = 135 to 180 g/t . Reagen ts of second cond i t i one r : H2SO 4 = and c leaner 2: sil icate = 270 to 315 g/t .

59.0 13.9 46.4 59.6 9.O 35.9 59.3 11.8 82.3 34.5 64.3 6.5 77.4

= 360 to 450 g / t ; t a l l oil L s = 225 to 270 g/t ;

1035 to 1225 g/ t ; 1st rougher cell, c leaner 1

1977), constructed for the sulphonate process, was started by using fatty acids; the tests done during the construction being sufficiently attractive to justify the change.

The essential parameters of this flotation process are: (a) Desliming in the vicinity of 15 micrometres, which is indispensable.

This is carried out by separation in cyclones in two stages. (b) Conditioning in thick pulp, if high recoveries are to be obtained with a

minimum addition of collector. This is the key-phase in the process. It takes 5 to 8 minutes for a pulp with 65 to 70% solids.

(c) The most selective fat ty acid is oleic acid but the froth formed is difficult to operate. It is more attractive to use a tall oil composed of a mix- ture of oleic acid and linoleic acid. Moreover, the froth is controlled by the addition of a frother (MIBC).

(d) The temperature during conditioning and flotation, and the hardness of the water axe also important features. In winter, the results drop slightly because of the lower temperature, which provokes a greater hardness.

The most characteristic example of this technique is given by the plant of the "Republic Mine" (Johnson and Bjorne, 1964) which treats approximate- ly 24,000 t /d of a specular hematite ore which contains small quantities of magnetite and martite in a gangue of recrystallized chert in which sericite, grunerite, cummingtonite and chlorite occur.

The important points in the circuit are: (1) Desliming in two stages, the underflow of the cyclones feeds the

188

Crude Ore (-16 ram)

L.2Mi. Clas!ilier

0,6 mm

F;pir'al~ ~, Concentrate

~lt Tailing

Ball Mill

U/F Desliming O/F =- Slime l (2 stages) 1

Densityin¢ Thickener O/F ~ Thickener

l U/F 6 conditioners

Rougher flotation J C1-2 -----~C3-12 ~ TaLling

Recleaner flotation

C1 - 7

Concentrate

o/

Concenlrate

R apLfine screen

l u/s Elutriation

Concentrate

Crude Ore

Rod Mill

Scalpin~cyclones

l Ball Mill u/v 1

Classifying cyclones

O/F 1 ~ O/F Desliming E l u t r i a t o r -

U/F

Magnetic separators

I NM O/F DensiIying thickener 1

I U/F Flotation -------~q'ailing

Fig. 2. Groveland 1961.

Fig. 3. Groveland 1976 (line 4).

conditioners i and 6 (there are 8) where the tall oil is added (560 g/t approxi mately).

(2) According to the characteristics o f the ore, the circuit may constitute a rougher, a scavenger and two cleaner flotation stages or one rougher, two scavengers and one cleaner flotation stage. The exchange of the pipes allows for a rapid transformation of the scavenger bank to the cleaner flotation stage and vice versa.

(3) To improve the grade of the concentrates, and considering the regrinding required to attain the conditions for pelletization, a reconcentra- tion circuit was installed to benefit from the supplementary liberation of quartz and residual silicates. It was observed (Erck and Nummela, 1962) that a new conditioning stage at elevated temperature depressed the silicates present, without complementary addition of reagent. Therefore, steam injec-

189

tions (or a heat exchanger) were installed in the pulp conditioner at 70% solids. The temperature is progressively raised to the boiling point. The flota- tion which follows yields iron grades in the concentrate of over 66%.

Figure 4 and Table II illustrate the double flotation circuit. A more recent utilization of this type of conditioning in very hot pulp was

realized by H. Laapas {1975) on various magnetite or ilmenite ores. A condi- tioning between 80 and 100°C in the presence of activating Cu 2+ and acid (H2SO4 or HF) as a depressor of the gangue, allows a very great selectivity of the flotation of iron ores (with a fatty acid or a tall oil).

To alleviate the iron losses brought about by the desliming, a flotation method without a desliming stage was tested by Hanna Mining (Bunge and Pocrnich, 1960). It consists of a ground pulp of 70 to 90%, - 4 4 micro- metres, to accomplish a conditioning in thick pulp (65 to 70%) of a mixture of a fatty acid type Acintol FA -- 2900 g/t and of fuel no. 2 (2600 g/t) at natural pH for 20 to 30 minutes. Sodium silicate {approximately 600 g/t) is added to the flotation cells as a depressant. There are 3 or 4 washing stages.

Crude Ore

Rod MiLl

Classification

u/F 1 Ball Mill

i

Slimes ~ O/F Cyclones ~ O/F Cyclones

U/F [ U/F

I I I~[ I I I I,I

L Rougher flotation w,, 1 (or 2) scavanger flotation -Ta/llng i

1 (or 2) cleaner flotation

Primary concentrate ~, Filtration

Pperegrind sLlos

i Regrind bail mills (80 ~ - 44 ~ m)

½ Steam conditioners I 98°C - I O(PC ,i

I 1 st Clcnir fl°t at i°n - - ~ S . . . . ~ger OotaUon.

L--3rd cleaner flotation 2nd retreatrnent

[ 2nd Clca~r flotation--W,,~ 1st retieatment

1 I _ _ Final concentrate Tailing

Tickener DiscfUters Pellet plant

Fig. 4. R e p u b l i c M i n e s c h e m a t i c f l o w s h e e t .

190

TABLE II

Metallurgical balance of "Republic Mine"

Products % weight % Fe % Fe distribution

Flotation Concentrator Concentrator Flotation circuit feed feed circuit

Regrind-refloat concentrate 90.2 46.6 66.90 97.8 84.3 Refloat tail 9.8 5.1 13.83 2.2 1.8 Regrind-refloat feed 100.0 51.7 61.70 100.0 86.1 Primary concentrate 54.3 51.7 61.70 89.1 86.1 Primary float tail 45.7 43.4 8.92 10.8 10.3 Primary float feed 100.0 95.1 37.59 100.0 96.4 Secondary cyclone overflow 4.9 25.50 3.6 Concentrator feed 100.0 37.00 100.0

The present t rend (Abeidu, 1976; Yang, 1977) in the research on anionic direct flotation is to aim at the use of a mixture of sodium silicate and salt (ferric sulphate for Yang, ferrous or aluminium sulphate for Abeidu) as a selective agent. The pH is slightly alkaline (pH -~ 8.8) and the fatty acid is added in a quanti ty of 450 to 650 g/t. A frother (DF 250 or MIBC) or fuel regulates the quality of the froth. The conditioning is done in diluted pulp (20 to 25%) at a duration of approximately 8 minutes.

Another possibility of anionic flotation can be found in flotation in an acid environment. The plant at La Perla (Mexico) floats the iron oxides with oleic acid in a sulphuric environment (Anonymous, 1980).

PARTICULAR ANIONIC FLOTATION METHODS

Certain ores, like that of Carol Lake, contain gangue components of the carbonate type in addition to silicates. The preceding anionic procedure therefore has to be adjusted to obtain purified ferriferous concentrates.

It was possible to improve on two processes, without, however, arriving at an industrial application.

(1) Double flotation, using a petroleum sulphonate (R 899) or a fat ty acid to float the carbonates in an alkaline circuit followed by a second flota- tion with R 899 in an acid circuit (pH = 3) or by a fat ty acid in a weakly acid circuit to float the iron oxides. Table III gives the results of a pilot plant using a fat ty acid in the two steps.

(2) Use o f various anionic reagents. The reagents of the hydroxamates family were able to provide possible uses of interest. These collectors are very selective with respect to carbonates, but they axe very sensitive to fines and extended desliming must precede the conditioning. The residual hydroxamates in the railings is of a toxic level that will disturb life in the rivers in which the water is disposed of. Table IV shows the possibilities of

TABLE III

Results o f pi lot plant - - Dual Process

191

Produc t s % weight % Fe % SiO2 % Fe d is t r ibut ion

Fine spiral tail 100.0 27.9 43.9 100.0 Slime (combined) 13.1 19.4 9.1 Carbonate concen t ra te 19.5 11.06 47.5 7.7 Iron concen t ra te 32.4 65.1 5.0 75.5 Iron tailings 35.0 6.1 7.7

Reagents Carbonate reagents, g/t I ron reagents , g/t

H2SO , - - 270 Fa t ty acid L5 170 275 Gum 9072 105 -- Fuel oil 90 - - NaOH 125 - - Sod ium silicate - - 245

TABLE IV

Locked cycle h y d r o x a m a t e f lo ta t ion results

P roduc t s % weight % Fe % SiO~ % CaO % MgO % Fe d is t r ibu t ion

Feed 100.0 15.5 100.0 Concen t ra te 14.2 62.2 1.7 1.9 1.7 56.9 Tailings 77.8 6.4 32.2 Slimes 8.0 21.1 10.9

separation with 100 g/t of an alkyldimethylammonium hydroxamate (active at 75%) associated with a fatty acid (18 g/t) and fuel (360 g/t).

FLOTATION OF SILICA BY AN ANIONIC REAGE N T

From the beginning of the sixties, the anionic flotation of silicates through the activation of silica by calcium ions was tested by several labora- tories, in particular those of Hanna Mining and the U.S. Bureau of Mines, in the form of discontinuous tests as well as p~lot circuits (at 900 kg/h).

This type of flotation implies (Frommer et al., 1964; Bunge et al., 1977) a depression of the ferriferous minerals which are collected at the bo t tom of the cell. The reagents which are used most commonly are gums, various starches in the form of gels or causticized, and dextrines. The silica is floated in a basic environment by a fat ty acid after activation by a calcium salt, usually calcium chloride.

This type of flotation can be done without desliming, notwitb-*anding the

192

fine grinding (60 to 100%, - 4 4 micrometres). The conditioning is done in stages in pulp concentrates ranging from 25 to 64% at a total duration of approximately 4 minutes. The required pH to activate the silica is in the vicinity of 11.5. Lime can replace calcium chloride. It must be remarked that siliceous particles larger than 74 micrometres are difficult to float.

Table V and Fig. 5 summarize the results obtained during the pilot plant test of semi-taconites.

The costs of the reagents is responsible for the fact that this method has not been developed in the manner which the results would suggest.

TABLE V

Reverse anionic flotation

Products % weight % Fe % SiO~ % Fe distribution

Concentrate 58.4 60.3 6.0 90.5 Tailings (froth) 41.6 8,9 9.5 Crude 100.0 38.9 100.0

Reagents, g/t Point of addition

NaOH = 1600 1st conditioner Gum 9072 = 1100 1st condi t ioner CaC12 = 765 2nd conditioner Fatty acid (Acintol FA2) = 720 2nd conditioner

Gum NaOH

CaCI 2 FA 2

Cyc l°°'ek~_J

O/F I U/F

~ . 1st conditioner

=. 2nd conditioner

Thickener I st cleaner flotation S

F i l ter 2nd cleaner flotation S

Concentrate 3rd cleaner flotation

Tailing

(h"oths)

Fig. 5. Schematic flowsheet of anionic silica float pilot plant.

193

F L O T A T I O N OF SILICA BY A CATIONIC R E A G E N T

The principle of the method is similar to the previous one. After depres- sion of the ferriferous minerals by reagents of the starch or dextrine family, the silica is collected by a cationic reagent normally without requiring ac- tivation by calcium ions, until at basic pH, the silica and the silicates have a negative surface charge.

The collectors are part of a large group of amines R-NH2. At first used in the form of acetate or chlorhydrate of alkylamine, these reagents were diffi- cult to administer, specifically at the level of production: difficult dissolu- tion once the alkyl chains were of the stearyl or oleyl type. Only the amines of the lauryl type showed a certain degree of ease of dissolution. The froth remained difficult to manipulate if the desliming was insufficient.

The appearance on the market of beta amines and ether amines has per- mitted of surmounting an important step. They proved to be more easily dis. persed and soluble, less sensitive to the variations of pH, more tolerant with respect to the presence of very fine particles -- at an identical cost price. Al- though the amines are more expensive than the fat ty acids and tall oil, the quantities required for the flotation are considerably lower, rendering the process very competitive. Almost all plants which adopted the flotation (partly or entirely) as a means of beneficiation, function by using cationic reagents (Jacobs et al., 1978).

They work {Tables VI and VII) either: (a) on crude ores: Sept Iles, Samarco for example; or (b) on gravimetric preconcentrates: Bong (and mag- netic); or (c) on magnetic preconcentrates: Empire, R~serve (Allie, 1980), Meramec. In the pilot stage, retreatment of magnetic or gravimetric railings has been considered (Houot and Polgaire, 1979).

The first patents referring to this type of reagents are already old (Harris, 1939, De Vaney, 1945, 1949). The most common reagents of the ether amine type were patented from 1968 onwards (Cronberg et al., 1968; Anonymous, 1977a). The combined use of alkyl and ether amine dates from 1976 (Polgaire, 1976; Houot et al., 1977).

The principal manufacturers of these products are: Ashland (USA) with the series of Arosurf {e.g. MG 98 A and MG 83 A); Hoechst (West Germany) with the series of Flotigam (e.g. ENA); and CECA S.A. (France) with the series of the Cataflot {e.g. C81).

Examples of reagents: MG 98 A = 3-n-nonoxypropylamine acetate MG 83 A = N-(n-tridecoxy-n-propyl)-l ,3 -- 'propylene diamine monoaceta te C81 = mixture of amino-l-alcane and an amino ether.

Certain reagents are more specialized with a view to the grain size of the products. The MG 98 A for instance is recommended for very fine pulps (100% of the product is - 4 4 micrometres) whereas the MG 83 shows good results for coarser particles.

194

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In certain instances, cationic flotation can be improved by using activators. It has been observed that the addition of calcium chloride in particular is favourable at doses which can attain 440 g/t (Dicks and Morrow, 1976; Houo t and Polgaire, 1979). The chlorides and sulphates of calcium, barium and magnesium in particular are more effective modifiers in the case of deslimed pulps. At identical recoveries, the iron content of the sink frac- tion may be improved by 1 t o 2%.

The iron oxides are depressed by starch and derivatives which seem to react in three ways (Polgaire, 1976): (a) by hydrophilization of surfaces (essential role of amylopectine); (b) by the more pronounced affinity for the iron oxides than for quar tz ;and (c) by the interaction of amylose with the collector, which increases with the number of nitrogen atoms in the chain and decreases with the rise of the number of atoms in the chain.

The use of amidons of amylose {e.g. of the corn type} is preferable to that of amidons which are solely amylopectinic (e.g. of the Waxy type).

Figure 6 and Tables VII and VIII give an example of the treatment at Sept Iles of ore from Schefferville {Canada). If one compared the flowsheet with that of Samarco (Brazil), one finds that the essential difference is the desliming stage, which is absent in Sept Iles and adopted in Samarco (Major-Marothy, 1972; Sleeman, 1976; Anonymous, 1975--1977b).

Crude Ore

I Pr, imaJ'y cascade Mill /

uIF J Cyclone

i O/F Primary Rapifine Screens

Bal! Mill

S

I 3 Cells scavenger circuit

'

Concentrate (to pellet plant)

u/s

u/s

Conditioners S]

8 Cells rougher Circuit F

4 Cells 1st cleaner circuit

J 4 Cells 2nd cleaner cwcmt

Tmling

Fig. 6. Schematic flowsheet of cationic silica (Sept Iles beneficiation plant).

This originates from mineral associations which are more or less enriched in limonitic and clayey fractions. The Schefferville ore is for a large part com- posed of massive silica ore, blue hematite and martite. Only the " red" frac- tion possesses a certain quanti ty of aluminium silicate; but this fraction only

196

TABLE VII

Reagents used in reverse cationic f lotat ion

Reagent, g/t Sept Iles Samarco Bong Reserve Empire

NaOH 900 225 -- H2SO 4 900 -- - - Starch 225 to 450 -- Dextrine WW82 = 450 -- 150 Amine MG83A = 90 MG98A = 67 to 112 ENA = 100 Frother A71 = 27 22 to 67

(13--15 C) = 90 5 MIBC = 45 11

TABLE VIII

Design and operational balance in Sept Iles (Canada)

Period Products % weight % Fe % SiO 2 % LOI % Fe distribution

Plant design Crude ore 100.0 56.0 13.7 3.5 100.0 Tailings 16.5 20.1 65.3 1.5 5.9 Concentrate 83.5 63.1 3.5 3.9 94.1

1975 Crude ore 100.0 55.6 16.5 -- 100.0 Concentrate 79.3 62.8 6.2 -- 89.6

1976 Crude ore 100.0 56.0 14.7 -- 100.0 (4th quarter) Concentrate 81.9 63.3 5.4 -- 91.6

r e p r e s e n t s 10% o f t h e c r u d e o r e a f t e r h o m o g e n i z a t i o n . O n t h e c o n t r a r y , t h e B r a z i l i a n o r e was s u b j e c t e d t o s t r o n g e r s u p e r g e n e a l t e r a t i o n , a n d is c o n t a m - i n a t e d b y f r a g m e n t s o f l a t e r i t i c c r u s t w h i c h a re n o x i o u s a n d a re r e a d i l y c r u s h e d t o a f i ne s ize .

In t h e case o f r e t r e a t m e n t o f t h e p r e c o n c e n t r a t e s b y f l o t a t i o n , d e s l i m i n g can o f t e n b e a v o i d e d , b e c a u s e t h e b u l k o f t h e p r i m a r y s l imes , w h i c h a re m o s t t r o u b l e s o m e , d i s a p p e a r d u r i n g t h i s p r e c o n c e n t r a t i o n s t ep .

OBTAINING OF SUPERCONCENTRATES

A t t h e p r e s e n t t h e r e is a c e r t a i n t r e n d in t h e p r o d u c t i o n o f s t ee l t h a t o r i e n t a t e s i t t o w a r d s a c h a n n e l d i f f e r e n t f r o m t h e t r a d i t i o n a l b l a s t f u r n a c e f o l l o w e d b y t h e c o n v e r t e r s t e p . I t is s p e c i f i c a l l y r e c o m m e n d e d in a r eas r i ch in gas, in w h i c h case t h e p r e r e d u c t i o n is f o l l o w e d b y t h e e l e c t r i c s t ee l f u r n a c e .

As a r e s u l t o f t h e s m a l l s p e c i f i c w e i g h t o f s lag ( 1 0 0 kg a p p r o x i m a t e l y ) c o m p a t i b l e w i t h t h e s e t e c h n i q u e s , t h i s n e w c h a n n e l r e q u i r e s c o n c e n t r a t e s a s s a y i n g less t h a n 2% s i l i ca ( o r 2 .5% SiO2 + Al~O3) w h e r e a s t h e t r a d i t i o n a l r o u t e ( i n t a k e o f 3 0 0 k g o r m o r e ) d o e s n o t i m p o s e s u c h c o n s t r a i n t s .

2/16 mm

O~F

Crude Oi,e (- 200 mm)

Rockcyl Mill

I 6/32 mm Siz ing

- 2 a m l "

Cyclone [

Pebble Mill

Symons cone

t I ~,. Bin

32/76 mm

Condd ioner -q

Desli~ung Th ickene r

Condit ioner ..~

Rougher flolation

• 4 ~ 1st c l e a n e r notation

' ~ - 2rid c l e a n e r flotalion

4 - - 3rd c leaner flotation

~ 4th c l e a n e r Ilotation

Tai l ing (frolh)

Corn S t a r ch

Slime(TaLling)

Corn S ta~'ch

Amine

Corn sta~ch

Concentrate

Fig. 7. Tilden Mine schematic flowsheet.

197

For the ores preconcentrated in magnetite (containing 4 to 5% SIO2), the U.S. Bureau of Mines developed the following method (Tippin, 1972; Veith, 1974; Colombo, 1977):

(a) Conditioning during 8 to 14 minutes in thick pulp with 1 to 2 kg/t caustic soda which is added during attrition.

(b) Desliming to eliminate the fines and to reduce the pH. This stage causes a negligible loss in iron (0.5% approximately).

(c) Flotation of silica by an ether amine (200 to 300 g/t). Concentrates of 70.2% iron assaying less than 2% SiO2 (1.3 to 2%) with recoveries which can reach 95% were obtained in a pilot circuit in which flotation was combined with magnetic separation and screening of the +44 micrometres fraction.

A mixed magnetite-hematite ore was preconcentrated by gravity (spirals). The ore, assaying 63.7% iron and 6% silica, was then subjected to flotation, Cataflot C81 (110 g/t) at a pH = 10.8, and 500 g/t o f causticized corn starch at 25% being used. It was possible to obtain superconcentrates of 68.9% and 0.9% SiO2 with an output of 87.6% (Houot et al., 1980).

198

The Meramec plant (U.S.A.), which can produce pellets with a SiO2 con- tent of less than 3%~ is an industrial example of production of superconcen- trates (Anonymous, 1976). A fraction of this concentrate is again improved by reverse cationic flotation to produce magnetite which is used in a heavy medium product containing 71.5% iron and 0.15% SiO2. The product is sold at the rate of 5400 to 8100 t /y for the manufacture of ceramic magnets.

SELECTIVE FLOCCULATION AND REVERSE FLOTATION

The very fine particles are usually detrimental to the efficiency of flota- tion of ores of the "ox ide" type. They consume large quantities of reagents (because of their great specific surface) and render the froth difficult to operate.

Consequently, when the beneficiation problem of oxidized taconite of Mesabi Range and hematite-goethite jaspilites of Marquette Range arose, the use of desliming cyclones appeared to be indispensable. But the loss of iron in the ultrafines was too significant. This loss incited the U.S. Bureau of Mines to devise a selective desliming process which allows of conserving the ferriferous fines, whereas the fine silicates are eliminated. Accordingly, as from 1962 onwards, research was undertaken on a process of selective flocculation. These studies were soon after crowned by success and certain results were published by Frommer from 1964 onwards, and a patent was issued in 1966 (Frommer and Colombo, 1966 ; Frommer, 1968).

The principal phases of this type of t reatment are: (1) Dispersion of particles coming from the crusher, obtained in an

environment in which the pH is maintained at approximately 10.5 by condi- tioning with soda and addition of dispersants (sodium silicate, tripoly- phosphate, chestnut extract, etc . . . . ) at moderate levels to avoid retarda- tion in the next phase.

(2) Addition of a flocculant and conditioning of pulp (10 to 20% in solids). The ferriferous particles agglomerate and are decanted, whereas the silicates remain in suspension and are evacuated by overflow. The flocculant that appears to be the most efficient is a starch (e.g. tapioca flour) which is causticized for a short moment at a temperature of 90°C to promote its dissolution).

(3) The decanted and thickened felTiferous product is then submitted to reverse anionic or cationic flotation (Tables IX and X).

The first results were obtained with a reverse anionic flotation. But be- cause of the world progress realized in cationic flotation, this route became the complement of the selective flocculation (Colombo and Frommer, 1976; Dicks and Morrow, 1978; Colombo, 1980).

The Russian laboratories (Gubin et al., 1975; Capkov, 1977) were also en- gaged in this type of flowsheet. The selective flocculation phase is done with a mixture (1.3 kg/t) of soda and lime (which permits of attaining a certain alkalinity in the circuit and the presence of Ca 2÷) associated with sodium

TABLE IX

Inf luence o f selective f loccula t ion on reverse anionic f lo ta t ion

199

Selective Products % weight % Fe % SiO~ % Fe dis t r ibut ion f loccula t ion

No Concent ra te 42.9 63.9 6.6 75.0 Tailings 53.5 14.7 21.5 Fines 3.6 35.4 3.5

Crude ore 100.0 36.5 100.0

Yes Concent ra te 46.5 65.1 4.8 83.7 Tailings 43.0 9.9 11.8 Fines 10.5 15.4 4.5

C:cude ore 100.0 36.2 100.0

Reagents, g/ t Without selective With selective f loccula t ion f locculat ion

NaOH 1.81 1.85 Sodium silicate -- 0.92 Tapioca flour 0.65 0.58 CaCi: 0.52 0.39 Fa t ty acid 1.17 0.55

TABLE X

Selective f loccula t ion - - cat ionic f lo ta t ion reagent

Products % weight % Fe % SiO 2 % Fe dis t r ibut ion

Concentra te 40.7 64.0 4.7 77.8 Tailings 39.9 12.2 14.6 Fines 19.4 13.1 7.6

Crude ore 100.0 33.5 100.0

Reagents Point of addi t ion Quant i ty g/t

Sod ium hydrox ide Sodium t r ipolyphosphate Sodium silicate Tapioca f lour Dextr ine 8079 Arosurf MG98A

Dextr ine 8079

Aerosurf MG98A

DF 250

Rod mill 1000 Rod mill 125 Rod mill 375 Condi t ioning prior to selective f lo ta t ion 125 Condi t ioning preceding rougher f lo ta t ion 375 Pump prior to rougher and first cell of the rougher bank cell 90 F ro th product f rom first rougher f ro th cleaner 250 First cell of 2-cell unit - - rougher under f low cleaner 90 Single-cell middlings cleaner f lo ta t ion

200

humate (1.25 kg/t) for 12 minutes. Because of the addition of a tall oil (0.25 kg/t), the latter flotation is an anionic reverse flotation of silica activated by Ca 2+"

This type of process has been applied commercially. The complex of Tilden {adjacent to Empire Mine) in the Marquette Range treats at the moment approximately 18 million t /y of an ore that contains 35 to 36% iron and 45 to 50% silica, the latter in the form of chert. The iron minerals are essentially martite associated with lesser amounts of hematite, magnetite and goethite. The grinding size which is imposed by the mineralization of the ore is between 10 and 35 micrometres if a concentrate containing 65% iron and 5% SiO2 is to be obtained. The crushing in two stages (compare flowsheet 7) allows of obtaining a product to be treated which comprises 85% in mass of particles below 25 micrometres (Sisselman, 1975; Paananen and Turcotte, 1980).

The reagents used are summarized in Table XI.

TABLE XI

Actual reagent consumption at Tilden

Reagent Reagent addition locations Quantities, g/t (one year average)

Design Actual

Sodium hydroxide Sodium silicate Corn starch Amine Polymer (total)

Lime

Primary mill 925 470 Primary mill 250 250 Deslime thickener feed and flotation 800 850 Rougher feed box and 6th rougher cell 135 110 Concentrate thickener feed, concentrate thickener overflow and filter feed 63 110 Deslime thickener overflow and flotation tailings 1500 1000

The flocculation step requires a sufficiently diluted (8 to 10% solids) pulp. The flotation conditioner is fed by the underf low of the desliming thickener which yields a product at 50% solids. Middlings of cleaner flotation froth products are then recycled to feed the rougher flotation cells, which is done with a pulp containing 25% solids.

The concentrates obtained contain up to 67.5% iron with a SiO2 content equal to 5% at the most. The recovery is close to 75%. The key factor, the selective flocculation, is sensitive to variations in the composit ion of the water used. In fact, more than 90% of the water is reclaimed and it may con- tain ions and organic noxious impurities (7200 m3/h are re-used in that plant).

Table XII shows the large influence of the temperature on the various parameters which are controlled permanently. These variations induce varia-

TABLE XII

Seasonal variations of reuse water characteristics at Tilden

201

Controlled parameters Average monthly high Average monthly low

Temperature, °C 23 (July) 4 (December) Solids in suspension, mg/1 749 180 pH 11.3 (January) 10.7 (July) Calcium hardness, rag/l, as CaCO 3 36 (February) 5 (July) Total hardness, mg/1, as CaCO 3 38 7 SiO 2 in solution, rag/1 237 191 Residual amidon, rag/1 20 (March) 2 (July) Residual amine, mg/1 3 (January) 1 (April to November)

TABLE XIII

Influence of seasons on the selective flocculation at Tilden

Products % weight % Fe % Fe distribution

Crude ore : Summer 100.0 34.8 100.0 Winter 100.0 34.5 100.0

Deslime thickener overflow: Design 20.0 12.5 7.0 Summer 28.7 14.2 11.7 Winter 25.7 12.5 9.3

Deslime thickener underflow: Design 80.0 41.8 93.0 Summer 71.3 43.1 88.3 Winter 74.3 42.1 90.7

t ions in the pe r fo rmance o f the f loccula t ion as is indicated by the da ta in Table XII.

Pilot studies o f water rec lamat ion (Co lombo and Jacobs , 1976) indicate tha t the clarif icat ion level, as measured by turb id i ty , is inversely related to the calcium level o f the water t reated. Water having a tu rb id i ty o f 2100 p p m SiO2 (15 p p m Ca 2+) yielded iron concent ra tes of less good qual i ty than a solut ion o f a tu rb id i ty o f 4750 ppm SiO2 (10 ppm Ca2+), a l though the iron recovery was no t af fec ted by the calcium level.

The SiO2 con t en t o f the iron concen t ra t e can increase f rom 4.8 to 7.5% be tween 10 and 15 ppm Ca 2+.

CONCLUSIONS

Flo ta t ion is a very eff icient means o f benef ic ia t ion o f iron ores and compet i t ive with the magnet ic routes. As long as the desired concent ra tes are

202

of a mediumgrade and the mineralization is not too fine and composed of magnetite or hematite in a clean siliceous gangue, the magnetic separation process can remove the iron either at low or high intensity. This process in- volves moderate operational costs, although the capital investment is clearly higher (Jacobs et al., 1975). But as soon as the mineralization is mixed (mag- netite + hematite) and/or becomes very fine, flotation becomes a more attractive process.

Moreover, the procedure allows of obtaining higher grades in the final con- centrates, increasing the recovery figure. The froth fraction yields a mineral mixture that is difficult to isolate by magnetism without losing out on recovery. In the difficult cases, like the ore at Tilden, where the grinding size is below 30 micrometres, it is the sole method which allows of obtaining concentrates of an adequate grade and an acceptable recovery.

In most of the plants the control of the processes and the performances is done on a continuous basis. The quanti ty of silica, the predominant gangue mineral, is determined every four minutes by neutron activation (for in- stance Sherman, Tilden). The determination reacts on the flotation param- eters through the intervention of a similar computer (regulation of the addi- tion of amine to the rougher flotation stage).

At Tilden in particular, the evolution of the treated ore is closely followed. The drill cuttings are tested in such a way that reactions with respect to flocculation are anticipated, and the composit ion of the mean crude ore is kept constant as far as possible by means of an extraction pro- gramme that is fed into the computer. Moreover, the physicochemical parameters of the recycled water are studied continuously and the reagents adjusted accordingly.

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