PROCESS CONTROL AT THE ECSTALL CONCENTRATOR. Chief Metallurgist, Metals Division, Texas Gulf Sulphur Company, Toronto, Ontario.

Transcription

1 Chapter 30 PROCESS CONTROL AT THE ECSTALL CONCENTRATOR by R. J. C. Tait, P. R. Clarke, M. P. Amsden Chief Metallurgist, Metals Division, Texas Gulf Sulphur Company, Toronto, Ontario.. Vice President of Production, Ecstall Mining Limited, Timmins, Ontario. Mill Superintendent, Ecstall Mining Limited, Timmins, Ontario. INTRODUCTION Process Control at the Ecstall Concentrator begins in the open pit at the Kidd Creek Mine. While this paper, naturally, emphasizes the instrumental and technical features of the Process Control system within the concentrator itself, it would not be complete without some mention of the selective mining methods and grade control practices at the mine, and of certain features of plant design, which are simple, but highly effecti~e~elements of the Process Control system. The success of the entire Process Control system, from pit to concentrate load-out, owes much to the high degree of co-operation and mutual understanding among all the operating departments at Ecstall Mining Limited and to the willingness of management to accept, from the start, an overall control philosophy based upon the inter-dependence of mining and metallurgical operations. INFLUENCES OF MINERALOGY AND METALLURGY ON MINE PLANNING AND PLANT DESIGN Mineralogically, the Kidd Creek ores are fine grained and mostly very complex. Because of complex mineral associations and extensive inter-locking, very fine grinding and regrinding of middling fractions are necessary and selective flotation requires a very delicate touch. Four separate concentrates are made for shipment to copper, lead and zinc smelters. The Process Control system is tailored to the metallurgical objectives of maximum recoveries of copper, lead, zinc and silver into concentrates of optimum quality, for maximum economic returns. For example, zinc and lead in copper con-

2 centrates are not paid for, but, in fact, incur penalties beyond certain limits. Similarly, there is no payment for copper or lead in zinc concentrates and very little for silver. Because of the complex mineralogy, some metal values are inevitably locked into unsuitable concentrates but, by careful process control, selectivity can be maintained to the highest degree and such losses kept to a minimum. The task of the metallurgist is simplified considerably by natural segregation of three major ore types within the Kidd Creek ore body. Geographically, there are two distinct ore masses, referred to as the north and south ore bodies, both mined from the same open pit. The larger north ore body contains two major types of ore, a large section of relatively clean copper ore on the hanging wall side and, on the footwall, a major deposit of highly pyritic zinc ore containing lead and silver. The south ore body is made up of mixed ores of copper ard zinc, of varying grades and degrees of complexity, constituting the third major type. The mineralogical and metallurgical characteristics of the three basic types are, briefly, as follows: 1) North Copper Ore consists basically of chalcopyrite in a rhyolite gangue, with some pyrite but practically no zinc. Most of the chalcopyrite is finely disseminated throughout the gangue and to some extent associated with pyrite, but some occurs in high grade to massive stringers and patches. Grade is, therefore, somewhat erratic. This ore is very hard and requires very fine grinding but copper recovery, in high grade concentrates, is otherwise not particularly difficult. It represents about 20 percent of the total ore in the mine. 2) Copper Zinc Ores are much more complex and can be further classified as; (a) High Grade Zinc Ore consisting of very fine-grained sphalerite containing up to 1.5 percent copper in very complex association with the sphalerite. (b) Banded Copper-Zinc Ores of medium grade containing approximately equal contents of copper and zinc in intimate association. (c) High Grade Chalcopyrite Ore occurring in irregular masses in contact with the other types. Although often massive in appearance, this copper ore usually contains appreciable zinc in complex association with the chalcopyrite. Mixed ores comprise approximately 40 percent of the total ore body. Owing to mineral inter-locking and pre-activation of sphale-

3 rite, rejection of zinc from copper concentrates is difficult. 3) Lead and Silver Bearing Zinc Ores. These are confined to the footwall side of the north ore body and represent approximately 40 percent of the total ore in the mine. They contain substantial proportions of the total zinc and silver, and practically all of the lead in the mine. They are highly pyritic, with major sphalerite, minor galena, very minor copper and appreciable silver (mainly occurring as very fine grained native silver), in intimate association with pyrite, in varying degrees of complexity. Some sections are graphitic as well as pyritic, appreciable pyrrhotite and some marcasite occur in certain areas and metallurgy has been further complicated in some sections by deposition of secondary copper minerals (e.g., covellite) at grain boundaries and along cleavage planes of sphalerite and pyrite. In the initial stages of metallurgical investigation of drill cores, serious attempts were made to treat all of the major ore types together as one ore, permitting the cheapest possible open pit mining. This soon proved impractical for the following reasons: 1) The highly pyritic silver bearing zinc ores are the most difficult to treat by flotation. Their very important silver content is difficult to exploit. Fairly good recoveries can be obtained in low grade, highly pyritic concentrates containing 12 to 18 percent lead, with a high zinc content, acceptable to lead smelters because of their high silver values. When these ores were mixed with copper and copperzinc ores, copper recovery was noticeably inhibited, and satisfactory silver recoveries were only obtained in dirty bulk concentrates unattractive to copper smelters. All attempts at further cleaning, dezincing or copper-lead separation resulted in excessive silver losses. 2) Depression of zinc during copper flotation of the mixed copper-zinc ores was a serious problem. If treating these ores by themselves for extended periods, the metallurgist would continually be faced with the choice of accepting low copper recoveries in copper concentrates of zinc content acceptable to the smelter; or paying heavy zinc penalties, at some risk of outright refusal of shipments by the smelter. However, by blending the clean copper ore with the mixed ores, in approximately ore reserve proportions, good copper recoveries could be obtained and zinc penalties kept to a minimum. Wide variation in ore grades within each category of ore presented another problem. Indiscriminate shovel operation in the copper and mixed ore sections of the pit, for instance, could result in sudden variations of copper heads from 1 to 5 percent copper and in

4 zinc heads from 5 to 15 percent zinc. The same applied to the silver-lead-zinc ore. Lead heads could vary from 0.5 to 3 percent and zinc heads from 8 to 18 percent. Variations of this order were encountered frequently in large scale pilot plant work carried out in two mills in the Porcupine district, and metallurgical control during periods of excessively high heads was very difficult. All of these considerations led to an early decision to design the plant to treat two distinct types of ore, to be supplied by selective mining and, grade-controlled blending of the three major ore types in the pit. Mining practi-ces adopted for metallurgical control are: 1) Selective mining of the silver-lead-zinc ores and the hanging wall copper ores in the north ore body. 2) Where practicable, simultaneous removal and blending of north zone copper ore and south zone copper-zinc ores, in proportions specified by the grade control department. 3) Continuing grade control studies and planning of mining operations to ensure reasonably constant mill heads and permit fairly accurate weekly forecasts of the grades of ore to be treated, as well as long range forecasts. The grade of ore in each pit blast is first estimated by extrapolation from the values indicated by the closest of the original exploration drill holes, and by the grade of the ore established from the same area on the preceding bench. In mining, each block of ore is drilled off well ahead of blasting and the grade is finally established by assaying of samples of all blast hole drill cuttings. 4) After blasting, ore may be drawn simultaneously from a number of sources of known grade and characteristics, following the instructions of the grade control engineers, who specify the number of truckloads per day required from each to deliver their predicted grade and tonnage to the concentrator. 5) The system includes strict scheduling of the two ore types through the primary crusher, and storage in two separate 6,000 ton storage bins. Ore is hauled in separate trainloads to the concentrator receiving bin and each load crushed separately into the appropriate fine ore bin. These measures require a high degree of co-operation between the grade control department and the mining staff in mine planning, and the best of communications at all levels. Shovel operators, haulage truck drivers, primary crusher operatars, train crews and concentrator personnel all share the responsibility for the success of the operation. It is to the credit of all concerned that the system has been highly successful.

5 The following features of plant design were incorporated to permit separate treatment of two ore types and augment grade control measures. 1) For the designed treatment of 9,000 tons of ore per day, the plant was laid out in three, parallel, 3,000 tons, sections, with identical fine ore bins, and grinding circuits. At the west end of the mill, "A" section has flotation circuits designed to treat only the copper-zinc ores. Similarly, "C" section at the east end is designed to treat only the pyritic lead-silver-zinc ores, while "B" section in the centre can treat either type of ore. Principal differences are in numbers of flotation cells, since the pyritic zinc ore is much slower floating, and requires more extensive cleaning for optimum concentrate grades. 2) Two separate 100 foot thickeners, filtering systems, concentrate dryers, and load-out bins are installed for zinc concentrates. This permits separate sale of high-silver, zinc concentrates from the high-silver zinc ores to smelters which pay for silver. Zinc concentrates from the copper-zinc ores are low in silver and are sold under different contracts. 3) It was recognized that mine-run ores could not be blended very effectively. Although some mixing could be expected in coarse crushing and storage, the grade of ore hauled could still be very erratic. To prevent wide swings,in grade during milling, the fine ore bins are specially designed for blending. Each milling section has a fine ore bin of approximately 8,000 tons live capacity. During crushing, the bins are bedded by a tripper, which traverses back and forth continually over the full length of the bin. Ore is withdrawn from each bin by seven slot feeders operating in unison. The combination of bedding and continuous withdrawal through multiple draw points is most effective. Grade may vary from day to day, but swings in grade during the course of a day are negligible, and size segregation is minimal. These operating procedures and design features contribute very substantially to the overall success of the Process Control system. Grade control is not perfect. The ore components for the most desirable blend are not always accessible, but predictions are reasonably accurate. Extensive metallurgical tests have been done on drill cuttings from blasts all over the ore body. Knowing the sources of ore each week as well as the predicted grades and with prior knowledge of the behaviour of ores from the various zones, the metallurgical staff can anticipate problems. This is extremely valuable. PROCESS DESCRIPTION

6 The process consists of fine crushing, very fine grinding, selective flotation of copper concentrates, silver-lead concentrates, and two grades of zinc concentrates. In all cases, middling fractions consisting of cleaner tailings and scavenger concentrates from approximately the last half of each rougher bank are reground in concentrate regrind mills. In the treatment of copper-zinc ores the reground middlings from the copper circuit are returned to the head of the rougher banks. In silver-lead flotation and in all three zinc circuits; they are scavenged, in open circuit, in separate cell banks, and the regrind scavenger concentrates are sent to cleaning with the rougher concentrates. Each grinding circuit consists of a 10% ft. by 16 ft. rod mill followed by a 12 ft. by 18 ft. primary ball mill in closed circuit with 15 in. Krebs cyclones and a 12 ft. by 18 ft. secondary ball mill in closed circuit with 10 in. Krebs cyclones. Each circuit contains two 8 ft. by 12 ft. concentrate regrind mills in closed circuit with 6 in. Krebs cyclones. All mills were supplied by Canadian Allis Chalmers Ltd. In copper flotation, primary cycloae overflow, at 55 percent minus 325 mesh, is treated in three parallel banks of five, 60 in. by 66 in. Wemco Fagergren flotation cells to recover coarse chalcopyrite. Primary copper tailings are returned to the secondary ball mill circuit and secondary cyclone overflow at 80 percent passing 325 mesh is sent to secondary copper flotation. Copper concentrates are cleaned in three stages in No. 24 Denver flotation machines. Copper flotation is carried out at a ph of approximately 6.5 maintained by lime addition in the presence of sulphur dioxide as a zinc depressant. Reagent 208 is the principal copper collector, assisted by a lesser amount of Sodium Isobutyl Xanthate. Methyl Isobutyl Carbinol has proven to be the most suitable frother. Secondary copper tailings are conditioned with copper sulphate and sphalerite is floated with Sodium Isobutyl Xanthate, with lime as a pyrite depressant. Frother additions are rarely required, but small amounts of Dowfroth 250 are used on occasion. Zinc concentrates in the copper-zinc circuits receive two stages of cleaning. The treatment of the pyritic zinc ore is similar except that the primary flotation stage has been eliminated and secondary cyclone overflow is sent to silver-lead flotation in a soda ash circuit with sulphur dioxide as zinc depressant, Reagents are; otherwise, the same as for copper flotation. Zinc flotation is the same, except that three stages of cleaning are used. Except for copper cleaners, all cells are 60 in. by 66 in. Wernco Fagergren machines. A11 concentrates are thickened, filtered and dried to approximately

7 6 percent moisture for shipment to smelters in Canada, The United States, Europe and Japan. CENTRAL CONTROL A Central Control complex, conveniently located between the grinding and flotation sections, houses in the main control room, the shift foreman's office, the rotameter room, the x-ray analyzer room and an instrument repair shop. With the exception of the concentrate drying section, all functions of the concentrator are operated and co-ordinated from the main control room, containing the essential control panels, the operator's desk, x-ray readout typewriters, the remote controls for the mill water pumps, and the Honeywell 21 and 316 computers and their auxiliary equipment. The control room, x-ray room, rotameter room, and instrument repair shop are equipped with their own air conditioning systems. Communications throughout the plant are relayed to and from Central Control via a public address system, direct line telephone and dial telephones. A VHF radio system links Central Control with the minesite, railroad and mobile units. Large windows on each side of the control room permit the operator a wide view of the grinding and flotation sections. There are five control panels in the Central Control room; one for fine crushing, three for grinding and flotation (one for each circuit) and one for services. All the operating panels are Honeywell-fabricated graphic panels with the flowsheets laid out in detail. All major items of equipment are depicted by models equipped with red and green running lights to indicate stop-start condition. Flashing red lights indicate electrical faults. The fine crushing panel is shown in Fig. 1. It is equipped with manual stop-start controls for crushing plant conveyors, screens, hydrocone crusher drive motors, oil pump motors, and hydroset pump motors. It also has tripper controls, tripper position indicator and control switches for dust collectors in the tripper gallery, transfer tower and dumping station. The panel has alarm lights for high-low levels in different surge bins and for blocked feed chutes to different conveyors and discharge chutes of the hydrocone crushers. It has a set of level lights for the three fine ore bins, indicating 10, 50 and 80 percent loading, sensed by Ohmart nuclear gauges. There are annunciator alarms for crusher lube oil and crusher motor winding temperatures, and ammeters for drive motors of the principal conveyors and crusher drive motors. Six T.V. screens are mounted on the panel to monitor selected points in the crushing plant. The panels for the grinding and flotation sections, Fig. 2, have stop-start controls for:

8 Figure 1 - Fine Crushing Panel Figure 2 - Grinding and Flotation Panel

9 1) Seven fine ore slot feeders. 2) Two collecting and one rod mill feed conveyor. 3) Mill drive motors, mill lube oil pump motor, mill air clutch for each rod mill, primary ball mill, secondary ball mill and regrind mills - a total of five mills in all. 4) Flotation conditioners and distributors. 5) Controls for banks of flotation cells. 6) Controls for tailings agitators, filter feed agitators and concentrate thickener drives. 7) Two controls for M.G. set, variable speed drive on the fine ore feeders. 8) Control for the dust collector in the fine ore feeder area. In addition, each panel has 31 ammeters recording loading on all cyclone feed, middlings,tailings and concentrate pumps, 29 annunciator alarm-drops to monitor bearing temperatures on mill motors, lubricating oil pressure, mill motor winding amperatures and low air on drive clutches for the mills. There are annunciator drops for rod mill feed water, concentrate thickener drive torque and filter operation. Also included are tonnage totalizers for rod mill feed and instruments for ph and density recording at critical process points. The service control panel, (Fig. 3) monitors the following ser- Figure 3 - Service Panel

10 vices : 1) Water, air and gas service. 2) Reagent distributi.on. 3) Power distribution. 4) H.V. and D.C. maintenance alarms. The water, air and gas service section of the panel has pressure alarms on the compressed air and natural gas systems, pressure on gland water, high and low level alarms for the water storage and diesel fuel oil tanks. This section also has running lights and ammeters for the different water pumps. The reagent distribution section has level alarms for storage and head tanks for eleven different reagents, with five spare alarms. The power section has temperature alarms for all transformers and lights indicating power outages and phase conditions of the primary transformer bank. The H.V.-D.C. section has 19 alarms on unit air heaters and makeup air heaters. All operating equipment with the exception of pumps, can be started from Central Control. Pump starting switches are activated by a gland water pressure switch and must be started locally. After a reasonably short power failure, all three grinding and flotation sections can be started up in about 15 minutes. CRUSHING The crushing plant in the concentrator is completely automated and can be operated by one man from Central Control. By design, it is part of the general mill building, facilitating supervision and maintenance. The principal elements of the plant are less than two hundred feet from Central Control, and the instrument panel is conveniently housed there in a dust-free, temperature controlled atmosphere with relatively short cable runs to the crushing plant, obviating the need for a.separate crusher control room. The control panel (Fig. 1) is laid out with television screens on the top and push-buttons along the bottom. The flowsheet starts from the left hand side with the train dumping ore, and ends on the right at the fine ore bins receiving the final crushed product. In the crushing process, a nominally minus six in. ore, from the gyratory crusher at the minesite, is reduced to minus 5/8 in. by one 7 ft. Allis Chalmers Hydrocone Secondary Crusher and two 7 ft. tertiary Hydrocones, at the rate of approximately 800 tons per operating hour. The circuit is closed so that all of the finished product has passed through Tyler screens, with 5/8 in. rod decks. Ore is delivered by trains of 18, 100 ton, railway cars and

11 dumped into a 2,500 ton receiving bin. Six vibrating feeders under the bin feed, via two pick-up conveyors and a long inclined conveyor, to a 50 ton surge bin, discharged by a vibrating feeder, onto a vibrating grizzly with 1-% in. slots. The grizzly oversize passes into the secondary crusher, and the undersize drops directly onto a rod deck screen beneath the grizzly. The secondary crusher product drops directly onto a parallel screen and both discharge to common oversize and undersize conveyors. The oversize conveyor transfers to a tertiary feed conveyor leading back to a 100 ton, recycle surge bin a- head of the two tertiary crushers. From the recycle bin, the ore splits evenly into the two tertiary crushers, each of which discharges onto its own rod deck screen. The tertiary screens are mounted in line with the secondary screens and all four discharge to the same two product conveyors. The finished product is conveyed to an external transfer tower whence it is fed back to the tripper floor above the fine ore bins. The circuit is centrally controlled as follows: 1) To avoid accidental mixing of ores, train dumping is regulated by the Central Control operator. A red light at the tracks warns the train crew that the opposite type ore ore is still being crushed and the train must not unload until the receiving bin is empty and the circuit has been completely cleared. 2) The plant is started by remote control after a one minute alarm activated from the panel. By following a set sequence, all equipment is started against the electric interlock. 3) The vibrating feeders under the receiving bin are started and set to deliver a feed rate of approximately 800 tons per hour, governed by a weightometer on the inclined conveyor, and indicated and recorded by a strip chart on the panel. The feeders have variable speed controls, but once the feed rate is set, little adjustment is necessary. They are protected by nuclear level controls in the chutes above them, so that they stop automatically, with pans loaded and two feet of ore in the chutes. Besides protecting the feeders, this prevents loss of heat caused by the dust collection system and ensures smooth starting and stopping of the equipment. The inclined conveyor is monitored by television so that belt tracking, ore loading, ore condition (lumpy, frozen etc.) and conspicuous oversize, tramp metal or wood in the ore can be observed. Nuclear level instruments monitor the depth of the ore in the surge bin at the head of the plant. At 40 percent full, the vibrating feeders automatically stop, and at 95 percent the main conveyor is stopped. A variable speed control on the vibrating feeder between the

12 surge bin and grizzly can be automated with the crusher kilowatts. This feature has not been used, and we see no immediate need for it. All screen and belt chutes are equipped with nuclear probes to monitor blockage. If plugging is imminent, an orange light on the particular chute model on the graphic panel flashes and an alarm sounds. Nuclear probes in the tertiary recycle bin indicate bin depths at 20, 50 and 80 percent full. At 80 percent, secondary crushing is stopped automatically through interlock until the load in the tertiary circuit has been crushed out. All belts up to and in the transfer tower are television monitored. The tripper is television monitored through a single T.V. screen and three cameras, one for each bin, activated by a selector switch on the panel. The tripper is automatic and controlled from the panel. It can be set over any of the three fine ore bins, and will then bed that bin automatically. The tripper chutes are alarmed against plugging, and a sonar device on the chutes permits untended bedding of the bins, to preset heights. Seven nuclear probes in each bin activate lights on graphic bins on the panel to indicate bin levels 10, 50 and 80 percent fullness and give a rough profile of the ore in the bins. Centralized control, with maximum application of instrumentation, has resulted in efficient operation with maximum protection of equipment. The entire crushing plant can be operated from the panel by one man but, because of the size of the plant, three operators police the circuit, and can be directed very quickly to any trouble spot by the operator at the panel, using the public address system. With this arrangement, 10,000 to 11,000 tons of ore can easily be crushed in two 8 hour shifts, leaving one shift a day clear for maintenance. GRINDING Ore is withdrawn from each fine ore bin by seven slot feeders, discharging simultaneously onto collecting conveyors leading to the rod mill feed be1 t. The slot feeders are driven through an M.G. set, by variable speed D.C. motors. The feed rate to each rod mill is governed by a Merrick, type E, weightometer on the rod mill feed belt and manually controlled by setting an indicator-controller on the strip chart recorder on the control panel. This section also controls rod mill pulp density automatically through a, water to ore, ratio controller. A change of setting automatically changes

13 the speed of the seven slot feeders. If one feeder stops, the others automatically speed up to satisfy the tonnage demand. Rod mill feed chutes are equipped with tilt probes to indicate plug-ups. The tilt probe is interlocked with the rod mill feed be1 t. The tonnages ground are totallized by the Honeywell 21 computer and typed out every eight hours on a shift basis. At midnight, the total tonnages milled for the past twenty-four hours are typed out. Open circuit rod milling is followed by two stage ball milling, with cyclones for classification. Primary and secondary cyclone overflow densities are controlled by Accuray gamma gauges located on the discharge lines of the cyclone overflow pumps, and recorded continuously, as percent solids, on strip charts on the control panels. The percent solids is controlled automatically by setting the indicator-controller set point on the strip chart recorder. The gamma gauges regulate the density of the cyclone overflow by controlling the water addition to the cyclone feed pump. Originally, gamma gauges were installed on all ball mill discharge launders for automatic control of the pulp density in each mill. This proved to be an impractical application because the launder restrictions where the gamma gauges were installed continually plugged up with coarse muck and ball chips and the gamma gauges were removed. Accuray gamma gauges, with Cesium 137 as their radiation source, were selected because they had an excellent amplification'factor (signal to noise ratio) as summed up in the Accuamp module. Water additions to the primary, secondary and regrind mills are controlled manually from the operating panels. The flows in g.p.m. are indicated by dial gauges. The power drawn by each grinding unit is shown by a kilowatt indicator dial gauge. A master kilowatt recording strip chart on each panel can be used to record the power drawn by any one of the grinding units, through a selector switch. All pump boxes in the grinding section are equipped with Warrick electrode probes to indicate high levels. The probes are connected to alarm lights and a horn on the Central Control panels. Temperatures of the grinding mill motor and trunnion bearings are monitored by the Honeywell 21 computer every 60 seconds. If the temperature of any unit should exceed a pre-determined limit, an alarm sounds on the operating panel and the computer types out the alarm point and the actual t e m p e r I n

14 has been taken and the temperature drops below the high limit, the computer types out, "temperature location 'x' is 'O.K."'. FLOTATION Most of the control elements in the grinding and flotation sections are designed to ensure steady state conditions, vital to such a sensitive metallurgical system. However, the sampling system incorporating continuous on-stream analysis and continuous ph control is probably the most important single metallurgical tool in the Process Control system. The validity of continuous on-stream analysis is completely dependent upon the validity of the samples delivered to the analyzer. After considerable study of methods in use, the system, shown schematically in (Fig. 4) was selected. It operates as follows: Figure 4 - Sampling System 1) Every pulp stream to be sampled for assay or ph control is passed, in its entirety, through a sampling conditioner, vigorously agitated by a powerful Lightnin mixer, designed for 100 percent suspension of solids in pulp. 2) Each of these tanks contains a Galigher glandless vertical sump pump, sized to deliver approximately 50 gallons per minute of homogenous slurry to a 2 ft. diameter by 3 ft, constant head, slurry holding tank, equipped with a suitable Lightnin mixer to maintain homogenous suspension, located in a sample room directly above the x-ray analyzer room.

15 3) The holding tanks overflow continuously through a battery of Denver automatic samplers which cut a sample every 20 minutes from each slurry stream. These cuts are continuously filtered in 12 in. by 12 in. Denver laboratory vacuum filters. Filter cakes are collected every 24 hours, providing a set of production composite samples for assay by wet chemical and fire assay methods. Official metallurgical production calculations are based on these composite assays. 4) Where required, ph is sensed continuously by electrodes mounted in the slurry holding tanks. 5) Continuous slurry streams of approximately 4 g.p.m. flow continuously to the x-ray analyzer, through 318 in. I.D. plastic lines connected to side outlets on the constant head slurry tanks. The x-ray analyzers are of the moving carriage type. Slurry streams flowing through, carriage-mounted, sample cells fitted with "Kapton" windows, are presented to stationary x-ray heads for radiation, on a timed cycle. Altogether fifteen slurry streams are analyzed (five for each milling circuit), namely; mill heads, copper or lead concentrates, copper or lead tailings, zinc concentrates, and zinc tailings. They are assayed for copper, lead, zinc, iron and silver and the results are typed out continuously on automatic typewriters located by the Central Control operator's desk, providing up-to-date intelligence for control of the flotation circuits. In detail, the system consists of two, Applied Research Laboratories Process Control X-Ray Quantometers, (PCXQ) Model 44000, with complete racking and controls for fifteen separate slurry streams each and with a Model Interface Console designed to read out all x-ray channels and provide digital and analog signals to a Honeywell 21 computer. Each unit has one tungsten target x-ray tube and six dispersive channel receivers to read Pb, Cu, Zn, Fe and Ag and pulp density information. Each unit is complete with twelve input and output lines with necessarycut-off valves designed to operate on a slurry flow of four gallons per minute. The units are equipped with SOLA voltage regulators that provide *I% at 4 KVA 1/60/230V output. Input to the regulator can vary *lo%. An ARL Model x-ray tube power supply is furnished, complete with control panel. The system has a control panel for the Model Console and a L. & N. Recorder that provides flexibility so that one x-ray unit can be operated in Manual Mode and the other one in computer Mode through the Computer Interface Console. The units are installed in an air conditioned room adjoining the Central Control room.

16 The computer is a Honeywell 21 computer consisting of a main frame containing: central processor with a programmer's console, real time control unit and "Magnetite" coupler. Some of the systems' design features are: core memory cycle time of 6 microseconds, word length - 18 bits plus memory and parity guard bits, core memory capacity - 16,384 words directly addressable, 16 different levels of priority interrupts, total analog inputs points and total digital input/output capability of 288 points. Peripheral computer equipment consists of the following: operator's console, four IBM Model B logging typewriters with 24 in. carriages and one inputloutput teletype Model 35 with 10 characters1 second paper tape punch and tape reader. A floating back-up power supply is capable of operating the computer system during minor power surges and total power loss up to 15 minutes. The x-ray analyzer system is set up so that one unit handles all the sample streams from the copper-zinc circuits and the other takes the streams from the lead-zinc circuit(s). Normally, the copperzinc system handles ten slurry streams and the lead-zinc system handles five slurry streams. Three stream positions in each x-ray rack are taken up by standardization briquettes prepared from samples of typical products. Figure 5 shows one x-ray rack and part Figure 5 - X-Ray Analyzer of the x-ray equipment. During a typical sampling cycle, the cell

17 is pushed into position in front of the cathode tube and the shutter on the tube opens and bombards the stream for 60 seconds with x-ray radiation. The secondary radiation emitting from the slurry is reflected back and caught by crystals positioned to interrupt the wave lengths for Cu, Pb, Zn, Fe, Ag and pulp density (scatter radiation). The secondary radiation energy is converted to a voltage reading from which the computer calculates a percent element content. The computer then types out the complete analysis of the stream on the typewriter representing the particular circuit from which the sample came. Each flotation section has its own typewriter. Under normal operating conditions, when two sections are treating copper-zinc ore and one section is treating lead-zinc ore, a complete analysis of each copper section plus recoveries is typed out every 12 minutes. The assays and recoveries of the lead-zinc sectionare typed out every 6 minutes. The computer is programmed to calculate the ratio of concentration of each concentrate and also the recovery of the major metal in each concentrate. At the end of each shift, the computer types out the average of all the assays for the last eight hours plus the recoveries based on the average assays. At midnight, the computer produces the average results for the day. An operator can also instruct the computer to type out the cumulative average for the shift at any time. The performance of the x-ray and computer equipment has been very good. Small plastic slurry agitators, immediately ahead of the x-ray units, were a continual nuisance, and the 318 in. plastic sample lines were frequently plugged by tramp material. These troubles were eliminated by removing the plastic slurry tanks, piping the streams directly to the x-ray from the side outlets of the constant head tanks and installing stainless steel screens over the outlets. For a long time, due to accumulation of lime scale, the Kapton sample cell windows had to be changed every shift to ensure reliable assays. Installation of a high pressure flush water system to flush out the sample lines, through check valves, on a timed basis, operated by the carriage position of the x-ray sample rack, has reduced the frequency of window changes to twice a week. The longest sample line in the mill is close to 400 ft. long, and some static heads are as high as 80 ft. We have some problems in keeping the slurries moving through the long lines without sanding out. The addition of lime to the flotation circuits adds to pumping difficulties by forming scale inside the lines, but we have kept downtime to a minimum by completely replacing all pump parts during each scheduled monthly maintenance shutdown, and actual availability of each sample stream for x-ray analysis is 95 percent or better.

18 Other problems =re discovered during the calibration of the x- ray units immediately after installation. These were connected with the method of calculating an assay based on the x-ray data. Assay typeouts were meant to be derived from a three equation calculation as follows: Assay = M.P.Q. M = pulp density correction factor = al+blx+clx2 where X = pulp density measured by scatter radiation. A set of co-efficients al, bl, cl were derived for each element in each stream. P = matrix correction factor = a2+b2mfc2m2 where m = concentration of interferring elements. A set of coefficients a2, b2, cg were to be derived for each element and applied to all streams. 11 I1 m and "P" correction factors were designed to be less than one. Q = conversion of voltage intensity to assay = a3+b3~+c3~2 where I = measured voltage intensity. A set of co-efficients a3, b3, c3 were to be derived and applied to all streams. As illustrated below, (Fig. 6) the co-efficients for all three equations were to be obtained from experimental data: Figure 6 A similar scheme was provided for the calculation of true pulp densities from scatter radiation measurements. It was found that equation "Q" required at least three sets of

19 co-efficients for each element - one each for the head, concentrate and tailing streams. Pulp densities, as measured by scatter radiation, were found to vary with the concentrations of certain elements, e.g., Cu, Fe, Zn. It did not appear that these relationships were consistent for every slurry stream. Since only one set of matrix correction co-efficients was provided for each element, it was not possible to obtain valid pulp density readings. Therefore, correction equation "M" could not be applied. Correction equation "P" was not calibrated because of the difficulty in interpreting matrix errors in intensity readings, without adequate pulp density control or correction. It also appeared likely that more than one set of correction co-efficients was required for each element. As a result of these findings, equations "M" and "P" are not currently in use. The first two terms of equation "Q" are being used to convert voltage intensities, with co-efficients a1 and bl (from the density correction) replacing the original co-efficients a3 and b3, to provide more flexibility amongst the various streams. In addition, pulp density typeouts are not being corrected for matrix effects. Variations in pulp density had the greatest effect on the x-ray assays of the concentrates and for the determination of silver. We have solved the problem of the concentrate assay by controlling the slurry density to the x-ray with gamma gauges. As yet, we have not solved the problem of obtaining meaningful silver assays. The effect of matrix is continually present whenever the ore changes. Our present method of correcting this is to adjust the co-efficients each time there is an ore change. Currently, this is being done by plotting the x-ray assays versus the chemical assays, on a seven day moving average basis, to detect major matrix changes. Our ultimate aim is to derive a method for continuous matrix correction using a computer. In spite of these problems, we have shown that the x-ray can give very accurate results (based on chemical assays as a standard) when the correct co-efficients to suit the type of ore are applied. We have established the following error tolerance limits for key assay typeouts. Copper Ores -% Cu % Zn - Lead Ores % Pb - % Zn - Head * Cu/Pb Conc. 25*0.50 5* f k0.50 Cu/Pb Tail 0.10* f f 0.50 Zn Conc f 0.50 Zn Tail 0.50f f 0.20

20 NOTE: Percentages are related to the respective actual assays. Conversion co-efficients for voltage intensities are ajjusted as these limits are exceeded. The assay typeouts have proven to be an invaluable tool to the operating staff for controlling the metallurgy of the three flotation sections. The 24 hour composite samples taken by the Denver samplers are used to determine the relative accuracy of the x-ray readings, and for production calculations, but we rely solely on the x-ray assays for metallurgical control. All flotation reagents, except lime, are fed to the circuits from the rotameter room adjacent to the Central Control room. Brooks rotameters are in - line surface mounted on separate panels for each division and on one test panel. They are glass tube, variable area, industrial type, with a needle valve for flow control and are provided with individual lucite wedge illuminators. Metering tubes are of borosilicate glass and all piping is of 316 stainless steel with teflon seals. These rotameters are extremely easy to dismantle and clean. Reagents are supplied from constant head tanks located above the rotameter room. A typical reagent panel is shown in (Fig. 7). Figure 7 - Typical Rotameter Panel All reagents, most of them in bulk quantity, are received, stored, mixed and held in holding tanks in the reagent area of the mill. The mixed reagents are then circulated continuously between the holding tanks and the rotameter head tanks. All tanks are equipped with high-low level alarms.

21 ph control of the flotation circuits is set up for automatic control from a pressurized lime lo~p. However, the original type of ph instrument employing an antimony ring reference electrode could not be made to operate satisfactorily. Poor flow characteristics of the pulp through the instrument made it prone to sanding-up, and to continual problems with the deposition of "lime" scale on the calomel reference electrode. We are currently having very good success with a Cambridge EIL Industrial ph meter model These instruments are immersed in the appropriate tanks in the sample head tank room. In the flotation circuits all roughing, and zinc and lead cleaning~ are done in banks of 60 in. by 66 in. Wemco Fagergren cells. The levels in these banks are automatically controlled by Moore Instrument Company level controllers. The level is regulated by a simple dart plug in the sandhole of each bank. The operator can adjust the levels for any desired setting within certain limits by adjusting a dial on the instrument which regulates the air pressure in a bubble tube immersed in the pulp. All pump boxes throughout the flotation circuits are equipped with Warrick electrode probes for high level indication. The probes are connected to alarm lights and a horn on the Central Control operating panels. These instruments have worked very satisfactorily but we had to modify them by insulating the probes so that they would not become short circuited by material bridging the probes. The next step in our continuous search to improve metallurgical efficiency is the application of computer controlled reagent adjustment. We have prepared a program, very empirical in nature, which closely follows the steps taken by an operator in making key reagent adjustments to suit flotation circuit metallurgy. The program makes use of the assays and recoveries derived from the Honeywell 21 computer. This information is transferred to a 4K Honeywell 316 computer which compares the data to targets the metallurgist has set, and depending upon the situation, sends out a signal to increase or decrease a reagent. We have installed Worthington, linear flow, precision valves in the lines feeding the rotameters. They will be calibrated to open or close "x" cc/min upon receipt of a digital signal from the computer. At the time of writingsthis system was being prepared for start-up. CONCENTRATE DEWATERING SECTION Flotation concentrates are thickened, filtered and dried to six percent moisture for shipment. Torque controls on the thickeners are set so that when the torque exceeds a high limit, the rakes will automatically raise until the torque drops below the set point. The thickener torques are recorded on strip charts on the Central

22 Control panel and high limits are indicated by panel alarms. Thickener underflows are discharged by gravity to holding tanks. The density of the underflow is controlled automatically by a gamma gauge and Clarkson valve assembly. The underflow controls are regulated from Central Control and the percent solids of the slurry is recorded on a strip chart. The thickener underflow is pumped from the holding tanks to the filters. Levels in the holding tanks were originally monitored by a standard bubbler level control system, which requires density correction and if the density is not kept within pre-determined limits, the system becomes very unreliable. We are currently replacing it with one using electrical conductivity principles which is independent of density. Thickened concentrates are pumped to the filter floor for filtering and drying. The filtering and drying section is separated from the flotation section by a wall and is controlled from a dryer control room by one operator. Two operating panels are located in this control room. One contains stop-start controls for the thickeners, filter motors and vacuum pump motors, and stop-start controls for dryer feed conveyors and dryer product conveyors to concentrate load-out bins, as well as ammeters for conveyor drive motors. Dryer throughput is regulated by Merrick Type L weightometers located on the dryer discharge conveyors and recorded on tons per hour stripcharts. The second panel contains the temperature strip charts, furnace drive ammeters, scrubber and exhaust gas controls, burner controls and panel alarms for the four dryers. Filtered concentrates are dried in Nichols-Herreshoff, multiple hearth, dryers consisting of nine hearths (zinc) and seven hearths (copper and lead). There are two dryers for handling zinc concentrates and one each for copper and lead concentrates. The dryers are heated by burning natural gas. Each zinc dryer handles 1,000 t.p.d. and the copper and lead dryers 750 t.p.d. each. Multiple hearth dryers are unusual in the base metal mining industry and the reasons for selecting them in preference to rotary dryers are as follows: 1) The concentrates could only be filtered to 15 to 17 percent moisture. These sloppy filter cakes would be extremely difficult to handle in rotary dryers. 2) The gas velocity in the hearth dryers is approximately one quarter of that in a rotary dryer, resulting in much lower dust burden.

23 3) The oxygen content of the furnace gases in the hearth dryer could be kept at a minimum by recycling scrubber gases, reducing the possibility of sintering the concentrates. 4) With fixed shells, the multiple hearth dryers were easily adaptable to instrumentation. 5) Product control with the hearth dryers promised to be very good. Dried concentrates would be granular and the moisture content could be controlled consistently within the narrow limits. 6) Operating costs of the hearth dryers would be less than for rotary dryers, although the capital costs would be higher. 7) The installation of the multiple hearth dryers required much less building space. Industrial building space in frigid northern climates is very expensive. Operating experience has shown that all expectations were met once initial start-up "bugs" had been corrected. The dryer operator controls dryer throughput, as indicated by the weightometer strip chart, by adjusting the speed of the filters and watching amperes drawn by the furnace drive. Moisture contents of the dryer feed and discharge are measured manually at regular intervals. Dried concentrates are conveyed to 1,000 ton capacity storage bins in a load-out building located over the railroad tracks. Concentrates are loaded automatically into rail cars by shuttle conveyors. The two zinc concentrates are loaded out on the south side of the building, while the copper and lead concentrates are loaded out on the north side. Each car is weighed as it is being filled on a Howe-Richardson Model 6553 track scale. Loading rate is about 300 tons per hour. SUPPORTING SERVICES The main supporting services for Process Control instruments are electrical and instrumentation. The electrical sectionis part of the Maintenance Department whereas the instrumentation section is in the Engineering Department. This gives instrumentation the autonomy they require and ensures that all instrumentation technicians are properly qualified by engineering standards. Originally, the instrument repair shop was located at one end of Central Control, so that the technicians would be close to their work, particularly the x-ray and computer, but because of dust problems encountered servicing pneumatic and electronic instruments in the same room, a new shop was recently built in a part of the mill, quite remote from Central Control, for the pneumatic servicing. This has worked out very well.

24 The mill instrument section is responsible for all instrumentation on the property, plus that at the minesite. A Chief Instrument Technician supervises a Mill Senior Technician and ten technicians. Each technician specializes in a particular function, such as computers, radio, television and nuclear devices, but by frequent job rotation, each man has the chance to learn all jobs. With this system, the work is done very efficiently as indicated by over 95 percent availability of all instruments. The shops are equipped with complete testing equipment to service all instruments including computer and x-ray equipment. THE VALUE OF THE SYSTEM Evaluation and economic justification of each element in the system would be an interesting exercise, but beyond the scope of this paper. In start-up alone, the advantages of good process instrumentation were very obvious. All too frequently, the initial start-up of a large, complex, base metal plant is an agonizing period of several months of effort to reach full design capacity at predicted metallurgy. In our case, although installation of instruments was far from complete, the most essential components for stability of operation were available, and the first 3,000 ton section was handling its designed tonnage, with reasonable metallurgy, within 48 hours. The next two circuits were up to tonnage and expected metallurgy within 10 hours. Admittedly, conditions for start-up were very auspicious, with a well designed plant, capable and experienced supervisors and pre-trained crews, but the role of Process Control was extremely important. The value of a fast startup in rapid return of capital investment is enormous. The continuous sampling and analysis systems were expensive to install, but have contributed very real dollar savings. The grade recovery curves for Ecstall ores are very steep. Aided by continuous analysis, the operators can control concentrate grades within 0.5 percent of metal content of the metallurgist's target, with rigid control of tailings losses. As a rough estimate, this has been worth at least $500, per year in improved copper recovery and zinc rejection from copper concentrates, ad of great value in the control of concentrate grades and recoveries in the treatment of the pyritic zinc ores. Visual estimation of the grade of the silver concentrate is virtually impossible and in the case of the pyritic zinc concentrate, very difficult. This metallurgy has improved significantly since the system became fully operational and the vanning plaques and binocular microscopes are gathering dust on the operating floor. In designing our control system, our objective was not automation, but rather to provide our people with truly effective tools for control of and optimization of metallurgy, and we have reached