THE HIBBING TACONITE COMPANY STORY 

B. w. Stunkard 

E. R. Tyler 

K. F. Jopke 

0. Borgeson 

c. L. Matson 

A. E. Englund 

G. N. Carlson 

R. A. Bauer 

Hibbing Taconite Co. 

Hibbing, Minnesota 

For presentation at the 1979 AIME Annual Meeting 

New Orleans, Louisiana, February 18-22, 1979 

PREPRINT 

NUMBER 

79-119 
HIBBING TACONITE COMPANY 

INTRODUCTION 

Hibbing Taconite Company is located by the cities of Hibbing and 

Chisholm near the center of the Mesabi Range, approximately 169 km (105 

miles) from the Canadian Border. The plant site is equidistant from 

these two communities, about 6.4 km {4 miles) from each. A 169 km (105 

mile) railroad connects us with the port of Superior, Wisconsin, where 

the pellet production is loaded into lake carriers for lower lakes 

destinations. 

Inunediately to the south of the mine area is the world famous HullRust-Mahoning Mine which produced more than 700 000 000 million tons of 

red ore between the years 1893 and 1973. 

During the period in which this country was being sustained by red 

ore mines~ such as the. Pickands Mather & co. managed Mahoning Mine, men 

were already thinking about future sources of raw materials for the 

steel mills. Beginning in the early 1930's, Pickands Mather, Bethlehem 

Steel and Youngstown Sheet and Tube Companies recognized that processing of 

the low grade iron ores, or namely the ferruginous silicate called taconite, 

would be required in order to provide for the future needs of the steel 

industry. Extensive research and exploration followed during the next 

decade. This work reve~l~d the magnitude of the available low grade ores 

and some of the requirements to develop these reserves. In 1941, the 

Minnesota State Legislature passed a taconite tax law which established 

that taconite would be taxed upon a production basis and thereby made it 

economically possible to acquire the extensive reserves for future 

operations. 

In 1942, a laboratory was established in Hibbing for the purpose of 

treating .taconite ores and to develop a flow sheet for a commercial 

operation. work in this laboratory led to the development of a pilot 

plant situated in the eastern end of the Mesabi Range which was ca.lled 

Erie Preliminary Taconite Plant. Work in this plant developed many 

refinements in the processing of the low grade taconite ores. It was in 

this plant that the commercial scale flow sheet for Erie Mining Company 

was developed. As a matter of reference, the time frame for this preliminary 

work was 1948 thru 1952. 

The next step was the announcement in 1953 of the Erie Commercial 

Plant which was completed in September, 1957, as a 7 600 000 MT (7.5 

million long tqns) pellet facility and later expanded to 10 800 000 MT 

(10.6 million long tons). 

With the knowledge and experience acquired during the construction 

and operation of Erie Mining Company, Pickands Mather participated in 

the development of the Hilton Mine at Shawville, Quebec, the Wabush 

Mines in Labrador and Quebec, the Griffith Mine in Ontario and the 

Savage River Mine in Tasmania, Australia. The last three of these 

properties were developed utilizing autogenous grinding in the flow 

sheet. Each of these pellet producing facilities contributed to the 

design and planning ~or the Hibbing project. 
- 2 -

Starting in the late 1960's and continuing into the early 1970's,, 

extensive feasibility and design studies were undertaken by Bethlehem 

Steel and Pickands Math.er toward developing the taconite reserves in the 

Hibbing area. In July 1973, plans to proceed with the projectwere 

announced by the two participants of the joint venture, which were at 

that time Bethlehem steef Corporation and Pickands Mather. 

The initial announcement at a press conference by Minnesota's 

Governor and the Presidents of Pickands Mather and Bethlehem Steel was 

for the development and construction of 5 500 000 MT (5.4 million long 

tons) per year pellet plant. 

Design of this new third.generation taconite facility was layed out 

to incorporate the largest and most efficient mining and plant equipment 

in use on the Mesabi Range today. In addition, a new and unique flow 

sheet for the plant was developed in which a single stage of crushing 

and single stage of autogenous grinding was to be used. In the indurating 

section, four-meter wide straight grate machines utilizing external 

coinbustion chambers capable o.f firing either coal or oil were to be 

installed. 

EvalQation.of the Hibbing Taconite Flow Sheet began in 1967. 

Testing to develop this flow she~t was conducted jointly by Pickands 

.Mather and Bethlehem Steel Corporation at the Pickands Mather Research 

Laboratory at Hibbing, Minnesota. 

Final design, engineering and construction for Hibbing Taconite 

Company was performed by the Bethlehem Steel Company Engineering Department 

in cooperation with Bethlehem's and Pickand~ Mather's Operating Department. 

During cpnstruction, the purchasing and procurement for Hibbing Taconite 

Company was the responsibility of the Bethlehem Purchasing Department. 

Construction began with land clearing and grading on September 1, 

1973. A new 6.4 km (4 mile) railroad spur was brought into the plant 

site by the Burlington Northern Railroad and the public utility company 

brought power into the plant site which was, until this time, an almost 

roadless forest area. All marketable timber was harvested by private 

contract loggers in order to salvage all useable trees. More than 

19 937 m3 (5 500 cords) of timber were taken ~rom Hibbing Taconite lands 

prior to the beginning of construction. 

Mine stripping was c01mnenced in 2.6 km2 (one square mile) area to 

the south of the plant and foundation construction for all major buildings 

was underway. During the course of construction, 63 subcontracting 

companies involving a total of 167 contracts were involved. Over 1,400 

workers were employed at the plant site at the peak of construction in 

April 1976. 

During construction in 1974 and early 1975, the owners of Hibbing 

Taconite dec:1-ded to expand construction to increase its size by 50 

percent and its production capacity from 5.5 to 8.2 million metric tons 

( 5 .4 to 8 .1 million long tons) per year. 
- 3 -

Pellet production at Hibbing Taconite began on August 28 of 1976. 

Our first train load of pellets left the plant on January 20, 1977 via 

Burlington Northern Railroad to the Lake Superior port of Superior, 

Wi~consin. During the interim, the pellets were stockpiled at the 

plantsite. 

The general layout of the project includes the main plant area 

containing the Mine Shop where mobile mine equipment is maintained, the 

Central Shops and Warehouse, Crusher, Concentrator and Pelletizer. To 

the south of .. the plant is located the Mining Area connected by the main 

haul road and to the north, the Reclaim Water Pump Station and Tailing 

B13.sin. Bordering on the west is the Burlington Northern Railroad and to 

the east, the Administration Building and the main entrance road~ 

The owners of Hibbing Taconite Company are Bethlehem Steel Corporation 

with 62.3%, R~public Steel Corporation at 16%, Pickands Mather at 15% 

and the Steel Company of Canada at 6.7%. Pickands Mather serves as 

manager of Hibbing Taconite Company. 

The total investment.in the project is $360 million. Employment 

currently at 1,160 will peak out at about 1,250 upon completion of the 

present expansion·. The operati.D.g level now at 5 500 000 MT (5.4 million 

long. tons) per year .will increase as the first of the new mill lines 

begins operation near the end of March 1979. 

~he year of 1978 ·was the first full year of operation, with production 

. slightly over the rated 5 500 000 MT (5.4 million long tons) per year 

capacity. With the completion of construction set for July of this 

year, expected productio~ will be 6 600 000 MT (6.5 million long tons). 
- 4 -

lUNING 

The Nesabi Range extends from west of Grand Rapids to the east of 

Babbitt, a distance of approximately 209 km {130 miles). The commonly 

used term range is derived from the line of high hills formed for the 

most part by the Giants Ridge granite. The Biwabik iron formation lies 

on the south slope and it dips at approximately 8 percent in a southeasterly 

direction. The Hibbing Taconite Company mine lies about mid-way between 

the two extremities of the Hesabi Range. 

Mineral leases, surface lands and permits required to make the 

Hibbing project viable were started in the early thirties and completed 

by 1970. 

In preparation for a long range m~n~ng estimate and preproduction 

mining plans, a series of diamond drill holes was completed in Section 

36. 

Also, it was decided that a bulk sampling program would be undertaken 

whereby representative fractions of the mineable taconite in Section 36 

would be obtained for test work in the Pickands Mather & Co. Research 

Laboratory. 

In order to obtain representative samples, it was determined that 

it would be necessary tg_make crude taconite ore available from four 

test pits. The trench locations were selected on the west side of 

Section 36 for three reasons. 

1. Close geological control of the sub-units from previous churn 

·drilling. 

2. An adjacent diamond drill hole section which could be used to 

correlate and compare the test pit and drill.hole analysis. 

3. Utilization of topography to minimize the stripping of surface 

and broken taconite. 

Two hundred fifty-four t-lT of magnetic taconite were taken from each 

of four test pits. From this composite, bulk sample laboratory tests 

were conducted which led to the plant flow sheet selection. 

In 1970, long range operating estimates and preproduction mining 

plans were formalized. 

Following the construction announcement in July, 1973, the 1970 

mining plans and estimates provided the details required to get the mine 

development underway. 

The Mahoning Mine which was managed and operated by Pickands 

Mather & Co. closed December 31, 1973. Hibbing Taconite Company acquired 

the equipment and facilities from Mahoning to start the Hibbing project. 

In April of 1974 1 land clearing began. 
- 5 -

Other than land clearing, the mine forces during the summer of 1974 

confJ.ned their activities to the overhaul of two 5m3 {6-1/2 cu. yd.) 

shovels; a fleet of five 65 ton trucks and eight 34 ton haulage trucks. 

The initial stripping began in the fall of 1974 using that equipment. 

The preproduction stripping called for the removal of 3 800 000 m3 

(5 000 000 cu. yds.)·of surface overburden on a portion of Section 36. 

A s_ection is one square mile, 2.6 k:m2 (640 acres) or 16 forties. The 

. surface stripping program uncovered taconite on 7 forties. 

Sub-grade taconite, defined as less than 10% magnetic iron content, 

scheduled for movement·in preproduction was limited to 380 000 m3 

(500 000 cu. yds.). This was the minimum volume necessary to meet the 

road construction program. Stripping of surface and, in particular, the 

sub-grade taconite prior to the plant operation was limited to reduce 

startup costs. 

·._Two mining cuts were developed prior to the plant operation. These 

2 cuts made 2 100 000 tons of taconite ore available to start.the plant 

operation. 

All preproduction stripping and development cuts were confined to 

an area to the south. and west of that fault to minimize metallurgical 

problems in the plant operation. 

Oxidation of the taconite along the fault is quite extensive. 

Prior to stripping, magnetometer surveys were run to set the surface 

stripping limits •. They were rerun after the surface was stripped and 

rotary drill hole samples were taken before the final alignment of the 

taconite development cut running parallel to the fault was established •. 

I~ was decided to build up maintenance forces in early 1975 and 

erect shovels, haulage trucks and drills with Hibbing Taconite Company 

personJ?,el to. familiarize: :them with the equipment to assure a trained 

work force. 

The erection. of ·shovels, haulage trucks of 170 ton capacity and 

rotary drills has·heen an ongoing project that nears completion at this 

date. 

The first drill was_moved from the erection site to the mine in 

August of .1975. The distance from the erection site to the mine development 

area was ~-4 km j4 miles). The trucks were driven as· they became operative. 

The.first 12.2 m (16 cu. yd.) shovel was walked from the erection site 

to 'the mine in October of 1975. 

Major equipment acquired for mine operations consisted of: 

l. Seven 12.2 m3 {16 cu. yd.) capacity electric shovels. 

2. One 7.6 m3 ·(10 cu. yd.) capacity electric shovel. 

3. Nineteen 17Q ton diesel electric driven haulage trucks. 
- 6 -

4. Four 120 ton electric wheel haulage trucks. 

5. Seven rotary drills. 

Auxiliary eqUipment consisted of: 

1. . Eight rUbber tired dozers. 

2. Eight crawler tractors. 

3. Four road graders. 

4. Two 70 ton truck mounted cranes. 

5. Two sprinkler trucks with an 30 283 liter (8000 gallon) 

capac;ity tank. 

6. Four front end loaders of 7.6 to 11.5 m

3 (10 to 15 cu. yd.)· 

capacity. 

7. A tire handling unit. 

~s previously stated, preproduction stripping of waste taconite was 

restricted ~o that amount required to meet the requirements of the road 

construction program. 

From the onset, construction of permanent main haulage roads was 

conducted as outlined by the following guidelines: 

If the original top of ground was void of muskeg, pit run surface 

was placed to designed grade elevation. As the road progressed, a •. 9 to 

1.2 m (3 to 4') lift of waste taconite was placed and 203 rnrn (8") of pit 

run gravel was·kept abreast of the waste taconite fill. After a period 

of time, as the voids in the rock lift filled, additional pit·run gravel 

would be placed to maintain the road elevations. At· this time 1 the road 

crown would be established. At a later date, sections of the road would 

be covered with a 102 to 152 mrn (4 to 6") lift of top dressing. 

If muskeg was encountered, it would be excavated, back filled with 

blocky waste taconite to former elevation of the muskeg that was excavatede 

Thereafter, construction would be as detailed. The permanent haulage 

roads are constructed' with a minimum width of 30. 5 m · {100 1 ) • The roads 

require constant grading, re-application of top dressing and re-establishment 

of the crown. However, the often observed rolls and swells of heavy 

duty truck haulage roads have not been experienced. 

Drilling the relatively soft taconite on the Central Mesabi Range 

with rotary drills ha~ presented no problems. Drilling rates and tons 

per foot of hole drilled have both exceeded expectations. 

For the year of 1978, drilling rates averaged 50 m (165') per 

shift. Bit life in the 311 rnrn (12-1/4") size averaged 549 m {1800'). 

The 381 rnrn {15") averaged 64{) m (2100') per bit. 
- 7 -

There are no significant ~vents to report in the drilling and 

blasting of taconite other than the problem of over fragrr~nting of the 

taconite ore. 

To elaborate on the last statement, the fully autogenous mills at 

Hibbing Taconite require a uniform blend from the mine for weight recovery, 

silica and size distribution. 

Size distribution is essential to obtain designed tonnage rates in 

the autogenous mills. It had been established that a feed whereby 40% 

is +152 nim (+6"} material, 40% is minus 76 mm (3") material and the 

remaining 20% between a minus 152 mm (6") and a plus 76 mm (3") size 

range provides maximum throughput. It is to this end the mining operation 

is geared when it comes to drill and blast pattern design. 

Pattern layouts are simple. 381 mm (15") rotary blast holes are up 

to a 11 by 11 m (36 x 36') burden and spacing. 311 mm (12-1/4") holes 

are from 7.3 by 7.3 m {24 by 24') and upward~ Initial detonation is 

usually at the north end. After the first four holes are pulled, the 

succeeding rows are controlled by 17 millisecond delays. With the tiein around the perimeter using 25 millisecond delays. The powder factor 

·in 1978 was .213 kg (.47 of a pound) of explosive per ton blasted. The 

powder :!;actor will he lowered in 1979. 

.. A 11.7 m (45') bla~~ hole on a 9.8 m (32') burden and 9.8 m (32') 

spacing would be loaded with 227 kg (~00 lbs.} of slurry and 431 kg {950 

lbs.) of anfo. Stemming would be a minimum of 6.1 m (20') and 38 mm (1-

1/2") crushed rock is used. The slurry is primed with two .45 kg (1 

lb.) primers and the anfo. with .45 kg (1 lb.) Individual down lines 

are used. 

The heavy bottom load of slurry is to pulverize the taconite and 

produce the minus 7.6 em (3") material requirement. The top, with 6.1 rn 

(20' or more) of stemming is to produce the +6 inch material for feed to 

the autogenous mills. Powder factors of less than .18 kg (.4 of a lb.) 

per ton are probably necessary to provide a better size distribution for 

the p~ant feed. This remains one of the mine operator's major objectives 

at this time. 

Prior to the detonation of the first blast at Hibbing Taconite . 

Company, a blast monitoring network was established. 

Eight monitoring stations are occupied for each blast. Eight sound 

level meters and two seismographs are used. Test blasts using two 45 

kg (1 lb.) cast boosters and an electric cap are the means used to 

provide the data as to go .or no go. The main blast is not shot unless a 

re~ording of less than 100 decibels on the linear scale is recorded at 

all 8 monitoring stations. 

As with most mine op~rators, problems have been experienced with 

some trucks of the 170 ton capacity, shovels of the 12.2 rn3 (16 cu. yd.) 

capacity and the rotary drills capable of drilling 311 mrn to 381 rnrn (12-

l/4 to 15") diameter holes. The problems with the drills were insignificant. 
- 8 -

The shovel and truck problems experienced have been corrected with 

the exception of th~ frame problem on some of the 170 ton haulage trucks. 

A frame modification repair program is underway at this time and are 

optimistic that the problem has been corrected. 

In 1978, 33.5 MT (33 million L T) of all material was moved as the 

operation produced the designed tonnage of 5.5 MT (5.4 million L T) of 

quality pellets. 

The Mining Department currently employs 325 persons in Operations. 

and 147 in Maintenance. During 1978, the production rate was 556 MT per 

man shift, excluding maintenance. 

In 1979, the mining operation is scheduled to handle 38 600 000 MT 

(38 million L T) of all material. 

Among mandatory permits secured for the project was one from the 

State of Minnesota that granted permission to pump makeup water required 

for the ~peratiori from any 1 or more of 5 abandoned red ore mines. 

Basically, these mines are a ground water source. Initially, two of the 

these mines located south of the mining area have been tapped to supply 

makeup water. A 508 mm (20") diameter pipeline runs to the plantsite 

from these two mines. 

Additionally, the permit provided for the construction of a tailing 

basin north of the plant: Only one small lake of approximately .16 km2 

(40 acres) was within the confines of the entire tailing basin area. 

Reclaim water is pumped from the basin. to the plant through two, 1.98 m 

{78") diameter pipelines. Of special mention is the fact that the 

entire system is closed as specified in the permits. 

Hibbing Taconite Company, by permit requirement, established 3 

stations to monitor water quantity on 3 streams north of the tailing 

basin. In addition, water quality is monitored. · This monitoring is on 

a weekly basis with reports required on weekly, monthly, quarterly and 

year basis dependent upon the particular requirement that is being 

monitored. 

While on the subject of monitoring, as required by the State of 

Minnesota, an ambient air monitoring network has been set up. It consists 

of 7 stations. The published EPA schedule is being followed. At this 

station, the·setup for monitoring the 24 hour EPA sample of suspended 

particulate can be observed. The value of the network, established in 

January of 1975, which was 18 months prior to plant operations, has 

already been justified. 
- 9 -

ORE DRESSING 

The Ore Dressing Department at Hibbing Taconite Company is responsible 

for the crushing bf 28 052 713 dry MT (28 200 000 dry long tons) per 

year of crude taconite ore and the subsequent grinding, magnetic separation, 

tailing disposal and delivery of concentrate to permit the production of 

8 200 000 dry MT (8.1 million DLT) per year of iron oxide pellets. 

The flow sheet represents the latest in taconite plant design and 

current metallurgical technology~ Its simplicity is unparalleled in the 

taconite industry. Bri·efly, it consists of only one stage of crushing 

and one stage of grinding. Each mill line has only one conveyor belt, a 

single autogenous mill, two slurry pumps, two stages of magnetic separation 

and one stage of cyclone classification. In addition, the tailing flow 

by gravity via a launder into a basin has eliminated the need for tailing 

pumps and thickeners • 

Test·pitting and initial pilot plant testing to develop the Ore 

Dressing flow sheet was conducted jointly by Pickands Mather and Bethlehem 

Steel Corporation at the Pickands Mather Research Laboratory at Hibbing, 

Minnesota. The owners of the Hibbing project recognized the potential 

that autogenous grinding offered due to the competency of the ore. A 

total of 53 pilot scale tests from early 1968 to mid 1970 led to single 

stage autogenous grinding that is described in this paper. 

CRUSHING AND STOCKPILING: 

The processing of crude ore all begins in the crushing and stockpiling 

system. One system serves 6 mill lines and the other serves 3 lines. 

TWo parallel 1.5 m (60") x 2.8 m (109") gyratory crushers are usually 

operated at a nominal 229 mrn (9") open side setting. Control of truck 

dumping, crushing and stockpiling is accomplished from a crusher central 

control room. 

In each crushing system, crushed crude ore drops into a 254 MT {250 

L T) surge bin beneath the crusher. The material from the surge bin is 

then conveyed by a 2.1 m (84") wide by 6.1 m (20') long two-speed apron 

feeder to a 1.5 m (60") wide steel cable primary ore conveyor belt. 

This·in turn transfers the crude through a moveable stone box or chute 

to a traversing 1.5 m (60") wide steel cable shuttle conveyor positioned 

nearly perpendicular to the primary conveyor. The shuttle conveyors 

distribute the crushed crude to a stockpile with a combined capacity of 

350 539 MT (345 000 L T) which is a four day supply of feed for the 

mills. Live capacity of the stockpile is approximately 61 000 MT 

60 000 L T). 

MILL FEED SYSTEM: 

In each mill line is a 1. 4 m (54") t..,ride by 5. 5 m (18') long variable 

speed apron feeder located under the center of the crude ore stockpile. 

Located near the toe of the pile is a 1.4 m (54") \...ide by 2.4 m (8') 

long vibratory feeder. The primary source of feed to the mills is the 

apron feeder. The vibratory feeder is used as a standby unit and as a 
- lO """' 

source of coarse grinding media from the edge of the stockpile when 

insufficient coarse material is being delivered to the crusher. The 

feeders deposit the mill feed on a 1.2 m (48") wide conveyor.belt which 

delivers it to the mill feed chute. Each conveyor is equipped with a 2 

idler-load cell - electronic scale which operates at a range of 0 to 762 

MTPH (750 LTPH). A wide range is necessary to record peak loads on the 

belt created by uneven loading by the apron feeder. 

AUTOGENOUS MILLS: 

There are nine autogenous mills each 11m (36') diameter by 4.6 m 

(15') long. Each of the mills is powered by two 6 084 metric h. p. 

(6000 h.p.) wound rotor motors utilizing liquid rheostats for controlled 

torque start. Six of the mil~s operate at 78% of critical speed or 

10.06 RP.t-1 ahd 3 operate at 74% of critical or 9.55 RPM. The mill is fed 

at a rate of approximately 40'6 MT (400 L T) per hour. A normal charge 

is 35% of total volume. Pulp density is presently manually controlled 

to achieve maximum mill throughput. Tests indicate the mill percent 

solids to be in the 65 to 70% range. Slurry is discharged from the mill 

through 9.53 rmn (3/8") slotted grates located at the outer periphery of 

the discharge head. Dischargers behind the discharge wear plates carry 

the slurry into a 4.1 m (13'-6") diameter by 4.4 m {14'-3 1/2") long 

attached cylindrical trammel. The trammel is equipped with rubber 

~creen panels wi.th 3. 2 It!ID_. (1/8") by 25. 4 mm (1") openings. The -3. 2 mm 

{-1/8") material drops into a rougher sump below the tronmiel. The +3.2 

mm (+1/8") is discharged into a bucket wheel at the end of the trammel. 

The bucket wheel elevates and deposits the material into an oversize 

return box, from which it is water jetted directly back into the mill 

through a rubber lined return tube. High pressure water is the transport 

vehicle. The amount of trammel oversize circulating load is unmeasurable. 

SEPARATION AND CLASSIFICATION: 

In the separation and classification circuit, the 3.2 mm (-1/8") 

trammel undersize collected in the rougher sump is pumped by a 457 mm 

(18") x 457 rom {18") steel pump to the rougher magnetic separator feed 

distributor. The pump is powered by a 406 metric h.p. (400 h.p.) motor. 

The rougher distributor divides the slurry to feed 14 single drum rougher 

magnetic separators in each mill line. 

Each of the counter current tank design rougher magnetic separators 

is equipped with a 914 mm (3') diameter by 3m (10') long drum. Field 

strength is 800 gauss at 51 mm (2"). The drum is driven at 21 RPM by a 

7.6 metric h.p. (7 1/2"· h.p.) gear motor through a chain and sprockets. 

Tailing from each of the machines flows by gravity into the main tailing 

launder .. The concentrate is laundered to the cyclone feed sump and pump. 

The 406 mm (16") x 356 mm (14") 608 metric h.p. (600 h.p.) variable speed 

steel cyclone pump delivers feed to a cluster of ten, 381 mm (15") 

diameter cyclones for classification. The coarse underflow is returned 

by gravity back to the feed end of the autogenous mill. The fine overflow 

fraction flows to the finisher distributor which divides it to feed 5 

triple drum concurrent tank finisher magnetic separators for final 

concentration. The finisher drums are the same size and powered similarily 

to the rougher separator drums. The strength of the finisher magnets is 
- 11 -

500 gauss. -Finisher tailing also flows by gravity to the main tailing 

launder. The final concentrate flows by gravity in a launder to the 

Pellet Plant.. Tailing from each mill line is collected in a trench 

lined with a high alumina brick. The trench discharges into a 290 m 

(950') long, 3m (10') diameter semi-circular wood .launder lined with 

rubb~r and sloped at 3/4-of 1%. 

RECLAIMING WATER: 

Tailing flows directly into a 28.3 million m

2 (7 000 acre) basin 

where process water is separated from the waste material and impounded 

behind dikes. This is a closed system as required under state permits. 

·water is ·taken into a reclaim pumphouse through two 2A m (96") diameter 

intake pipes •. The pumphouse contains ten, 914 mm (36") x 610 mm (24") 

horizontal centrifugal 3 549 metric h.p. (3500 h.p.) pumps rated at 

108 765 J;.PM (29 000 GPM) each~ Water is recycled to the plant at a 

norma,l rate of 757 000 LPM (200 ODO GPM) through two 2 m (78") diameter 

buried pipelines for a distance of 6.5 k.m (4 miles}. 

CONTROL: 

The .control of all mill ~ines is accomplished from a central control 

room. All mill equ~pment except the magnetic separators and a portion 

bf the cyclones i~ controlled from here. The 2 major control circuits 

govern mill power draw. ·and the classification loop. 

Mill power draw can be controlled in three separate modes which are 

manual, cascade or. computer. In the manual mode, the mill power draw 

signal is sent to the control room as indication only. The control room 

operator manually adjusts the apron feeder speed based on the indicated 

power signal. In the analog cascade mode, the power setpoint is set on 

the controller. The analog signal is sent to the power controller via 

the power sensor. The power controller and weight controller determine 

the feeder speed, based on the rate of deviation and the actual deviation 

from setpoint of the 2 process variables, which are power and weight. 

In computer control, the computer receives the mill power analog signal 

and converts it to a digital signal. The computer then operates on the 

signal via a pre-programmed set of operations. The resultant control 

signal is converted to analog and sent to the apron feeder. 

The computer controls power in a much tighter band which permits an 

increase in the power setpoint. 

The second major control circuit is the classification control loop 

around the cyclones. Density of the cyclone feed is the pr,imary control 

to achieve the desired silica level in the mill product. Classification 

control is based on the silica results obtained from concentrate samples 

which are gathered hi-hourly. Silica and -325 mesh grind results are 

relayed to the control room operator from the lab. The control room 

operator raises or lowers the cyclone feed density controller setpoint 

to stay within a silica range of 4.3 to 4.6%. A nuclear density gauge 

senses variation from setpoint and controls water addition to the cyclone 

feed sump to maintain the pre-set density. A variable speed cyclone 

feed pump maintains sump level through a level sensor. Variations in 
-.12 -

pump speed result in pres~ure changes in the cyclone header. The control 

room operator adds or subtracts cyclones to stay within a given pressure 

range. There are four remotely controlled cyclones per mill line. 

Control of grade has been good and has permitted meeting targeted specifications. 

METALLURGICAL AND OPERATING RESULTS~ 

In 1978 the plant average mill feed rates were 395 MT (389 L T} per 

hour as compared with a designed capacity of 406 MT (400 L T) per hour. 

Rates in excess of 406 MT (400 L T) were reached in 4 months dtiring the 

year •. Milling rates vary with changing ore characteristics and size. 

Swings as great as a 'minus 76 MTPH (75 LTPH) and plus 51 MTPH {50 LTPH) 

from rated capacity have been experienced. Early pilot testing and 

subsequent bulk sampling after plant startup indicated a properly sized·. 

crude f1=ed as essential to attaining grinding capacity. As previously 

noted, the desired blend contains 40% plus 152 mm (6") material, 20% 

minus 152 mm (6") plus .76 mm (3 .. ) ai::td 40% minus 76 mm (3 11). The primary 

crusher is also adjusted to avoid over-reduction of the crude ore. 

The autogenous mill grinding power averaged 23.o0 KWH per MT (23.38 

~VH/L T) in 1978. Since this includes inefficiencies through the 

pinion, gear reducer, mill motor and electrical system, it compares 

favorably with a pilot plant prediction of 21.4 KWH per ,MT (21.7 :K'WH/L T) 

at the mill shell. - - ' 

Mill availability was 83.6% in 1978. The 1979 target is 87%. Mill 

lining is responsible for the majority of mill doWn. time. Improvements . 

in this area will be discussed later in this paper. 

The percent magnetic iron in the crude ore averaged 21.48% and 

total iron 31.40%. 

The plant produced a concentrate containing 67.95% iron ar1d 4.57% 

silica. The silica target for 1978 was 4.70%. Careful blending of 

crude ore delivered to the crusheris.required to meet these specifications 

with a flow sheet that does not permit regrinding of off specification 

concentrate. 

The targeted silica content in the concentrator was reached at a 

fairly coarse grind level of 78.2% minus 325 mesh. The Blaine surface 

area of 1420 cm2 per gram provides high filtering capacity in the Pellet 

Plant. 

The average recirculating load from the cyclones to the mills was 

135%. This does not·include the mill trammel oversize fraction internally 

returned to the mill which was previously noted as unmeasurable. 

A steady improvement throughout the year resulted in an average 

magnetic iron recovery of 95.38%. An ongoing program between operations 

and staff metallurgists detects and corrects off target mag iron losses. 

Conversion of all rougher magnetic separators from double to sin.gle drum 

is now under way and will further improve magnetic ir.on recovery. 
- 13 ....; 

CRUSHER WEAR LINERS: 

The two major users of wear liners in ore dressing are the 1.5 m 

~60") crusher and the autogenous mills. Wear rates of the crusher 

mangan~se concaves and mantles have been exceptional to date. Rather 

than the 3.6 to 4.1 million MT (3.5 to 4 million L T) life orginally 

anticipated, we are experiencing an average life of 9.1 to 10.2 million 

MT (9 to 10 million L T) • Our most recent concave change was made after 

15.2 I'(lillion MT (15 mil:lion L T) • Consequently, only 3 concave changes 

were necessary since startup in 1976. Prime contributors to the extended 

life are a wide crusher setting and closely bedded and fractured taconite 

in the upper part.of the orebody. 

MILL WEAR LINERS: 

Mill liner maintenance is presently being done on a 15 shift basis 

with three liner crews 5 days per week. The original mill lining consisting 

of a plate-lifter bar design on the feed, shell anddischarge ends of 

the mill. soon after startup, it became apparent that it would be 

difficult to meet availability requirements with the large number of 

pieces necessary in a mill of this size. A testing program was started 

to investigate the use of fewer pieces and improving the wear qualities. 

Five mills have nowbeen converted to a one-piece single wave design of 

shell liners. The number of individual shell liners has been reduced by 

75% •. A similar reduction is anticipated on the feed end of the mills 

where testing will be done with the one piece design. 

on. the discharge end, the plate-lifter bar design has been retained. 

However,· larger plates and lifter bars now in use are giving longer · 

life. 

The original chrome moly plate-lifter .bars of 370 Brinnel hardness 

were soon replaced by 540 Brinnel chrome moly material. High chrome 

white irons of 600 Brinnel hardness have also been tested and shown to 

greatly improve wear life. Their use in other areas will be. evaluated. 

Te?tirig and evaluating of new designs and metallurgy will continue as we 

seek higher and higher mill availabilities. 

Our single stage autogenous mill steel consumption is considerably 

less than is being experienced in rod mill-ball mill circuits and semiautogenous mill-ball mill circuits. Total steel consumption rates are 

.about 15% of the semi-autogenous mill and rod mill-ball mill rates. The 

autogenous mill steel consumption rates should improve when the present 

standard'chrome moly liners are replaced with white irons. 

_PUMP WEAR LIFE: 

Pump impeller and shell life.has exceeded expectations. Since 

startup no rougher pump casings have beep changed out. Projected life 

of the casings is 25 000 hours. Rougher pump impellers now have an 

average life of 14 000 hours. Similar experience with the cyclone pumps 

shows 6000 hours average casing life and 4600 hours impeller life. 
- 14 -

MILL GEARS AND PINIONS: 

As with any new plant, some difficulties were experienced during 

startup. The major area in the mill was a concern about uneven load 

distribution across the gear and pinion tooth faces. This was corrected 

by a program of hand grinding. Extensive testing of tooth loading, mill 

and gear deflection, gear hardness, tooth straightness and tooth profile 

and spacing showed that all were within specified design limits. We are 

optimistic that gear and pinion problems have been resolved. 

CONCLUSION: 

In conclusion, the Ore Dressing flow sheet incorporates single 

stage crushing and single stage autogenous grinding with the largest 

autogenous mills in the world. Its simplicity permits good housekeeping 

practices, requires a low manpower level and lends itself to an effective 

preventive maintenance program. Continued improvements are being accomplished 

in areas such as wear material, mill throughput, power control and 

magnetic iron recovery. Major problem areas have been encountered and 

corrected. Experience to date has demonstrated that all design criteria 

can be met or exceeded. 
- 15 -

AGGLOMERATING DEPARTMENT 

The Agglomerating Department is responsible for the thickening, 

storage and blending of the high grade magnetic concentrate; dewatering 

.for moisture control on filters~ green ball preparation in the balling 

section and indurating to the desired physical quality on the travelling 

grate furnaces. The final product is conveyed to the loadout bins where 

it is sUbsequently loaded into unit trains. 

The flow sheet of the original 5 500 000 MT {5.4 million L T) plant 

is divided into four major sections. The first is thickening and 

dewatering wh~re the concentrate is thickened in one thickener, stored 

and blended in two slurry storage tanks. Haterial is withdrawn from the 

storage tanks with pumps and directed to eight filters for dewatering 

and conveying directly to the balling section or to an outside stockpile. 

The next major section is balling, where filter cake is stored in eight 

storage bins. There are two lines with four balling modules in each 

line. The green balls are conveyed to two travelling grate furnaces in 

the indurating section. The last major section is material handling 

where the final product is collected and conveyed to the pellet loadout 

bins, or diverted to an emergency stockpile. In the expanded plant, one 

additional· thickener and storage tank will be provided to complement 

four disc filters. One line of four modules in balling will prepare 

green balls for indurating on one additional furnace. 

In the thickening and dewatering section, the concentrate is received 

by gravity directly from the Ore Dressing Department at 45% solids in 

two 33.5 m (110'} diameter thickeners. 

The feed to the thickener is normally a 1420 cm

2

;gm Blaine surface 

and will average 22.3% ferrous iron. The thickener overflow is pumped 

to the concentrator where it is utilized as process make-up water. The 

concentrate is thickened to 70% and the underflow is pumped by rubber 

lined pumps to three slurry storage tanks. Each tank stores up to 4064 

dry metric tons (4000 DLT) of concentrate. An interesting feature .in 

the design of this section is the fact that the equipment is located 

outside and exposed to the harsh environment of the Northern Minnesota 

winters. 

Concentrate is withdrawn from the storage tanks and pumped to the 

filters through pressure distributors.· The 12 filters are 2.7 m (8' 10") 

in diameter, twelve disc, and operate at.610 rran (24"} of vacuum while 

producing a filter cake with 9.0% moisture. 

A series of conveyors normally transport the filter cake to the 

balling section. The last conveyor is 1. 8 m (72"} wide and equipped 

with air operated plows.· The plows are automatically controlled by load 

cells on the filter cake bins and divert the material to the bins preferentially. As an alternate, the filter cake can be stored outside. 

The material in stockpile can be reclaimed with a front end loader and a 

reclaim system which slurrifies the material and returns it to the 

concentrate thickener. 
- 16 -

In a typical balling circuit, filter cake is withdrawn from a surge 

bin, bentonite is introduced and the two products are mix~d in a rotary 

mixer. The mixed product is conveyed to a balling drum as new feed. 

The product from the drum is distributed to a vibrating screen. The 

seed pellet, or -9.5 rnrn (-3/8") screen undersize, is collected on a wide 

·conveyor and is recirculated as drum feed. A grizzly with 15.9 rnrn 

(5/8") opening produces a final product. The +15. 9 rnrn (+5/8u) material 

is discharged to a shredder and the shredder product is returned as drum 

feed. 

Material is withdrawn from a 76 MT (75 ton) filter cake bin with a 

variable speed table feeder. The bentonite feeder, drawing from a 15 MT 

(15 ton) bentonite bin, is a vertical, gravimetric, load cell type. The 

filter cake and bentonite is conveyed on a belt to rotary mixer. This 

conveyor is equipped with a process control scale which maintains the 

rate at which the filter cake is withdrawn from the bin and provides the 

signal to the bentonite-feeder to regulate the additive rate. Bentonite 

consumption rates average 8.9 kg/MT (20 lbs/L T) of pellets. Experience 

indicates a goal of 7.1 kg/MT (16 lbs/L T) is realistic. 

The mixer is a top feed, horizontal rotary type. The mixed product 

is combined with recycle from the seed pellet screen and conveyed to the 

balling drum. Each balling drum is 3.7 m (12') in diameter and 9.8 m 

(32') long, driven with a 100 h.p. motor. Each drum is equipped with a 

rotary cutter bar. 

The vibrating seed pellet screens in the original' plant are 2.4 m 

(8') wide by 6.1 m (20') long. 

The final product from four balling circuits is collected on a belt 

and conveyed to the indurating section. This conveyor is equipped with 

a dual idler, load cell type process control scale which provides a rate 

signal to the furnace control loop. The green balls discharge onto a 

1.4 m (54") wide oscillating conveyor, which distributes the product on 

a variable speed 4.4 m (172") wide belt. The wide belt provides an even 

feed distribution to the roller conveyor ahead of the pelletizing 

furnaces. Each roller conveyor assembly consists of twenty-three, 

152 rnrn (6") diameter chrome plated rolls. They are spaced with a 6.4 rnrn 

(l/4") gap and serve to re-roll the green pellets as well as rerttove any 

entrapped fines. The roller conveyor undersize is collected by_two 

transverse conveyors and conveyed back to the-filtering section where it 

is mixed with the new filter cake for reprocessing. The roller conveyor 

discharges green balls onto a reverse roll which provides a more permeable bed of green balls on the pelletizing grate. 

' 

There are three horizontal continuous travelling grate machines 

which are 4 meters (13') wide and 76 meters {249') long. Each furnace 

has 38 windboxes, 2 meters (6~') in length; 124 individual pallet cars. 

The total grate area is 304 square meters (3271 square feet) which 

results in a grate factor of 27.3 MT (26.9 L T) per square meter (square 

foot) per day at the designed production rate of 346 dry metric tons 

(346 DLT) per hour of fired product. With a 305 rnrn (12") green ball bed 

depth, the retention time is 32 minutes. 
- 17 -

Each furnace is divided into 7 zones. The green balls are initially 

. dried in the up-draft drying zone at a gas temperature of 315.6°C 

(600°F).. The gases leaving the top c:;>f the bed are normally saturated at 

82.2°C (180°F} and are removed from the process with the hood ·exhaust 

fan. · Further drying is accomplished in the downdraft drying zone at a 

gas temperature of 232.2°C (4500p). The drying process is followed by 

preheat starting at a temperature of 648. goc (1200°F) • In the preheat 

zone, the oxidation of the magnetite begins and the hood temperature is 

raised to prevent thermal shock and too rapid oxidation of the pellets 

on the top of the bed before entering the firing zone. 

The pellets are heated to the final induration temperature in the 

firing zones with downdraft "on gas" temperature of 1343.3°c (2450°F.) 

.This temperature ·is usually sufficient to produce a high strength abrasion resistant pellet. After-firing, where gas temperatures are normally 

l037.8°C (1900°F), is the zone where heat is recuperated from the top of 

the bed by lowering the on gas temperature and using the 1343.3°C 

(2450°F) pellets to heat the downdraft gases. The pellets are subsequently 

cooled to a final temperature of 93.3°C (200°F) in two zones. Cooling 

is accomplished by the. introduction of air at ambient temperatures by 

the cooling air fan. Each furnace is equipped with 5 process fans with 

a total connected metric h.p. of 15 464 (15 250 h.p.). The two waste 

gas fans, hood exhaust and windbox exhaust, are discharged to atmosphere 

thru 5664 A.C.M.M. (200 000 A.C.F.M.) wet process scrubbers. 

The heating of the pellets to 1343.3°C (2450°F) is accomplished by 

burning Bunker "C" oil in two external, side mounted, vertical combustion 

chambers which are top fired. Each chamber if 14.8 m (48'-6") high and 

4 m (13'-3") outside diameter. A 25.2 million Kcal/hour (100,000,000 

BTu/Hr.) burner is mounted .tn the top of each chamber which is the only 

source of external heat applied to the furnace. This furnace design 

will facilitate coal firing in the future. 

Another unique feature of the indurating section is the fact that 

all of the main process fans are located outside of the main building. 

This design reduced the initial capital cost, facilitates maintenance 

and minimizes the problems of heat, fumes and noise inherent with fans. 

The final product from each furnace is conveyed to a transfer 

·house •. At this point, a transfer bin preferentially provides approximately 30% recirculation of fired pellets for hearth layer that are 

conveyed back to a screening station in the main building. The 1.8 m 

(6') wide by 4.3.m (14') long hearth layer screens provide 100% +9.5 rom 

- .25.4 mm (+3/8" - 1") screened pellets for hearth layer to protect the 

· hardware on the furnace grates. 

Throughout the entire pellet handling system, dust collectors are 

provided to maintain clean in-plant conditions. 

From the transfer house, the final product is conveyed to the 

pellet loadout bins, each of which has a capacity of 12 193 MT (12 000 

long tons). In addition, a small 25 400 MT (85 000 L T) emergency 

stockpile facility is provided near the loadout bins. Unit trains of 
- 18 -

180-91 MT (90 L T} capacity cars are loaded uniformly as' the trains move 

under the bins. Average loading time is two hours. The trains are 

pulled a distance of 168.9 km (105 miles} to a transhipping terminal in 

Superior, Wisconsin. 

The entire pellet plant is monitored from a .central control room 

which is located in the indurating area. With the exception of the 

controls for individual main balling circuits and-the filters, the 

operator has control of the process 24 hours a day. 

OPERATING EXPERIENCE: 

Since the initial start-up in 1976, the Pellet Plant has experienced numerous highlights:in its short history. On June 21, 1976, the 

first filter cake was produced that was stockpiled. The flow sheet and 

equipment arrangement in the balling and indurating sections provide for 

a limited production of green balls from one drum which can also be 

stockpiled with the filter cake. This allowed training of operators in 

the critical filtering and-balling phases as well' as debug the equipment· 

prior to starting the first furnace. The first pellets were produced on 

August 28. 

Ten days later, when the furnace was on hold for·corrective work, a 

fire broke out in the area above the combustion chamber on the roof. 

What could have been a major disaster was confined to the immediate area 

above the No. 1 furnace with effective control by trained fire brigades. 

In the initial design, the fuel trains to the burners were located 

immediately above the hot chamber roof. It is believed that a leak from 

a pilot oil line directed oil on top of the hot chamber which·ignited 

the entire fuel train. As a result of this, the fuel train was relocated 

to a new control house which is an appendage to the main building and 

isolated from the chambers. The new piping arrangement to the burners 

on top of the combustion chambers was up-graded to heavier piping, 

welded fittings and high pressure flexible connections to eliminate the 

hazard of oil leaks. 

On the cool down of the No. 1 furnace, following the fire, refractory problems were observed that dictated a change in design and materials 

in the combustion chambers. Rebuild work was carried out while the fire 

damage was repaired. 

In the meantime, construction work on the No. 2 'furnace was completed and it was brought on line in October of 1976. 

Following repairs and modifica~ions, No. 1 furnace was returned to 

production in February df 1977. By June of that year, both furnaces 

were producing at rated capacity. In 1978, the first full year of un-

-interrupted production, tonnage slightly in excess of our 5 500 000 MT 

(5.4 million ton} design capacity was produced. 
- 19 -

DEWATERING SECTION: 

In the early operations of the filtering section, extreme wear was 

·experie~ced on filter bags, sector$ and the barrels. 

· In cooperation with the manufacturer, on a warranty basis, a significant modification program was undertaken to eliminate high wear rates. 

This work was completed in time to meet plant requirements. Original 

design. called for 102 DMTPH (100 DLTPH) and are now operating at 153 DMTPH 

(151 DLTPH) and producing a 9.0% moisture filter cake. 

· BALLING SECTION: 

.Some operating problems have occurred in the balling section. 

The original cutter bars were 406 rom (16"0) in diameter. Vibration 

was experienced with the 406 rom (16"0) bar and they failed. These bars 

were repaired and reinforced. The first failure was replaced with a 

large 572 mm (22 1/2") diameter bar. In addition to the larger diameter, 

the design has been changed to eliminate welding of the flutes to the 

main bar. The flutes are welded to a separate tube which is clamped to 

the main bar. This larger unit has been operating since April of 1978 

with excellent results. 

Mechanical problems~have been experienced with the 2.4 m {8') wide 

by 6 m (20') long vibrating screens. Failures were experienced of the 

main side body·plates, exciter mounting plate and cross strut members. 

Using a spare screen assembly, it was possible to keep ahead of the 

operating requirements by maintenance procedures developed to facilitate 

a complete changeout in less than 8 hours. As a result of these problems, 

roll screens are being.provided in the expanded plant. 

Improvern~nts have been achieved in our green ball s~z~ng program 

over the past 12 months. The +12.7 m -15.9 m (+1/2" -5/8") has been 

reduced from 44.5% to 35.3% and the +9.5 m -12.7 m (+3/8" -1/2") has 

been increased from 48% to 55.6%. 

INDURATING SECTION: 

Pioneering of the.large, external, side mounted combustion chambers 

is an important highlight in the indurating section. Earlier furnaces 

were equipped with multiple burner arrangements mounted in the side 

walls of the furnace hood which do not lend themselves to coal firing. 

The·original chambers were equipped with a large vertical burner 

located in the center of the roof. The brick in the side wall of the 

chambers was not supported. During the initial operation of the furnaces, 

hot spots developed in the wall and roof of the chambers. When the 

first furnace came down to ambient, an inspection revealed that the 

brick had slumped. The chambers were redesigned to include 2 water 

cooled annular support rings for the brick. The burners were relocated 

off center to achieve better control of the flame pattern in the chamber. Improved operating procedures have been adapted to ensure that the 
- 20 -

chambers never operate under positive pressure conditions. Since these 

changes, the operation of the chamber and refractory experience has b~en 

very satisfactory. 

Several significant performance factors in the indurating section 

were experienced in the first full operating year of 1978. During 1978·, 

fuei consumption averaged 75 000 Kcal per MT (306 000 BTU per ton). 

During the year, 4 months were operated at capacity with Kcal (BTU) 

consumption under the 74 400 (300 000) mark. 

Additional areas being considered for fuel reduction include: 

Improved green ball sizing which will provide increased void area 

in the bed. 

Reduced hearth layer on the grate and sidewalls. 

Higher production·rates made possible by an adequate grate factor 

design. 

Improved process control with an "on line" computer. 

In 1978, availability of the indurating section at 94.2% exceeded 

the design of 90% as a result of a totally committed preventative maintenance program. From tlie beginning, a scheduled maintenance program 

throughout the entire plant was followed. Each furnace is scheduled 

down every sixweeks for a 12 hour maintenance turn. Common equipment, 

or life-line systems such as hearth layer and filter cake units are · 

scheduled down every 4 weeks for 8 hours or less. The furnaces are 

taken to ambient temperature every 12 months for refractory repairs and 

major mechanical machine repairs. This major maintenance shutdown has 

. been accomplished in 5 days with minimal refractory repairs since startup. 

PELLET QUALITY: 

The pellet quality of the furnace product in 1978 is as follows: 

The +6.35 mm (+1/4") before tumble averaged 99.4% •. The +6.35 mm (+1/4"} 

after tumble was 95.5%. The minus 28 mesh after tumble averaged 3.9% . 

and the cold compressive strength was.210 kg (462.lbs.). Quality improvement goals are aimed at increasing the +6.35 mrn (+1/4") after-tumble and 

reducing the variance. 

Probably one of the most desirable features of the Hibbing pellet 

is the high reducibility at 82.6% as compared to the international goal 

of 60%. The percent swelling at 12.5% is below the maximum goal of 15%. 

The pellet sizing in 1978 was 35.3% +12.7 mrn -15.9 mm (+1/2" -5/8") 

and 55.6% +9.5 mm -12.7 mm (+3/8" -1/2"). Recent operating experience 

in the +9 •. 5 mm -12.7 mm (+3/8'' -1/2") range is running in excess of 70%. 
- 21 -

The chemistry of the pell·ets produced in 1978 was as follows: The 

total iron averaged 65.86%, the ferrous iron was 1.0% and the silica was 

4.94% on a dry basis.