Chapter 9

THE ELECT30LYTIC Z I H C PLAET OF RUHB-7, IIJK GbiBH .

Datte l f i , Vest Germany

H. R. Wuthrich, Hanager

A . von Ropenack, Super intendent

Abs t rac t

The M e t a l l g e s e l l s c h a f t AG decided i n l a t e 1965 t o b u i l d an E l e c t r o l y t i c Z inc P l a n t a t D a t t e l n (w. Germany), Lurgi-Chemie was e n t r u s t e d u i t h t h e engineer ing and e r e c t i o n of t h e e n t i r e p l a n t . Ground was broken i n February 1967 and t h e opera t ion was s t a r t e d on August l s t , 1968. The new E l e c t r o l y t i c Zinc P l a n t was designed t o produce a minimun of 100,000 m e t r i c tons of s p e c i a l h igh grade z i n c and 200,000 met r i c tons of su lphur ic ac id .

Concentrates from Meggen (2. Germany) and mainly T h i n s ( ~ a n a d a ) a r e roas ted i n two t u r b u l e n t l a y e r r o a s t e r s . The s u l f u r d iox ide con ta in ing gases a r e converted i n a "Bayer double c a t a l y s i s u n i t " t o s u l f u r i c ac id . The c a l c i n e i s leached ba tch wise, t h e l each r e s i d u e separa ted from t h e pregnant s o l u t i o n by counter-current th ickeners and drum f i l t e r s . F o r mainta ining t h e s ~ l f a t e balance a modified j a r o s i t e l each i s being used. A f t e r a two s t e p p u r i f i - c a t i o n t h e n e u t r a l s o l u t i o n i s mixed wi th s p e n t e l e c t r o l y t e and cooled i n unpacked atmospheric coo l ing towers. Three e l e c t r i c a l c i r c u i t s a r e supplying t h e power t o t h e tankhouse, each r a t e d a t 30,000 amperes. W i t h a maximum c u r r e n t d e n s i t y of 750 amperes p e r square meter and a s t r i p p i n g c y c l e of 24 hours , an average c u r r e n t e f f i c i e n c y of 90 pe rcen t i s obtained.

Cathodes a r e melted i n one induc t ion fu rnace and c a s t i n t o shapes, o r a l loyed wi th aluminum and copper f o r d ie -cas t metal .

The second purification residue i s treated for cadmium recovery in an electrolytic cadmium plant.

A total of 308 daily paid men and 59 employees a r e required to produce approximately 9, 500 tons of zinc. Maintenance labour i s not included in the foregoing.

Introduction

Metallgesellschaft AG has been active for many years in the production of nonferrous metals in Germany a s well a s in foreign countries. With the loss of i t s affiliates in foreign countries, af ter the world wars, a new approach had to be conceived. Moderni- zation became imperative particularly in the field of lead and zinc plants which were located in the western part of Germany. The old report plant at Berzelius was replaced by an Imperial Smelt- ing Furnace which successfully supplied lead and prime western zinc for the German market.

Simultaneously with starting up of the ISF operation, Metallge- sellschaft decided to expand further the zinc production by building an electrolytic zinc plant. In postwar Germany this was to be the f i rs t venture to produce special high grade zinc in one step and was possible because Metallgesellschaft was successful in obtaining electrical power ra tes at an economical level.

The reasons fo r the choice of Datteln, a small town on the north- east border of the Ruhr district, were:

I. Favourable transportation, the plant i s adjacent to the Rhein-Herne canal (inland waterway)

2. Consumers of sulphuric acid in the vicinity 3. Adequate labour supply 4. Reasonable power ra tes and a provision to

deliver excess steam from waste heat boilers to power plant.

Metallgesellschaft i s the owner of 140, 000 square mete rs of land. Of this amount 45, 000 square mete rs i s covered by plant buildings and operating equipment.

The land i s bound by the canal on one side, the federal railroad t racks on another and the main road leading through Datteln on the third side, forming a tr iangular shaped piece of property.

ELECTROLYTIC ZINC PLANT OF RUHR-ZINK GMBH 249

Fig. 1 - General view of the Datteln plant

Most foreign concentrates (shown by Table I ) a r e received by barges which a r e loaded at the sea ports of Rotterdam and Amster- dam. The barges a r e moved to Datteln along the Rhein River and Dortmund-Heme canal system. Domestic concentrates a r e shipped to the plant by Federal railroad. Most of the plant products a r e shipped to customers using the highway trucking system and by rail- road. The layout of the plant facilities conforms to the receiving and shipping facilities.

Engineering

By March 1966 the metallurgical flow sheet and optimum capacity was established. Lurgi Gesellschaft fur Chemie und Hiittenwesen mbH were instructed to design and construct the entire plant. In Feby. 1 9 6 7 preparation of the ground site was started and on August I st, 1968 the first concentrate was fed to the roaster (shown by Figure 2).

For the choice of equipment and the flowsheet the following factors were considered:

ELECTROI~YTIC ZINC PLANT OF RUHR-ZINK GMBH

T A B L E l a

R P i c a l Analyms o f different concentrates, Calcine and Leach Residue

T A B L E l b

TYpical Contents o f m i t i e s

c . 0 m o r n C1 F Co N i A s

Leach Residue 1 .8 4.76 0.9

Government permission would not be granted unless the emmission of gas and fume complied with the strict regulations and no discharge waste materials entering the biological sewage treatment plant would interfere with i t s normal operation.

A flowsheet and equipment were chosen which would enable the operation to comply with the government pollution requirements and be able to treat a variety of raw materials. At that time a leach residue treatment process was not included in the flow sheet. The choice of a residue treating process was delayed in order to study the various processes being developed. Plans t o include residue treatment were considered and space provided to add at a later date after the plant was in operation.

In order to obtain a maximum of flexibility it was decided to operate leach and purification a s a batch process. Fo r economical reasons a counter current thickener system for decantation of neu- t ra l leach solution and washing of leach residue was chosen. Fo r the final dewatering of the leach residue vacuum drum fil ters were provided.

For electrolyte cooling unpacked water cooling towers were specified, The necessity to evaporate a maximum of water made this type of equipment desirable.

The total employment i s shown by Table 7.

Concentrate Handling and Storage

The Meggen concentrate (shown by Table 1) which i s being pro- duced by a parentcompany in Germany, arr ives at Datteln in train loads of 750 - 1, 000 tons. After weighing and sampling, the bottom dump c a r s a r e unloaded into an underground hopper. From there a belt conveyor system transports the concentrate at a rate of 200 tons per hour to the covered concentrate storage building.

All other raw materials, at present mostly from Canada, reach Datteln over the German inland waterway. Barges up to 1, 375 tons holding capacity can pass the locks which a r e between the Rhein, the North Sea and Datteln. A harbor i s provided to accomodate four barges simultaneously.

The concentrate i s unloaded by crane into one of two automatic weighing hoppers which travel with the crane. The weighing system enables the operator t o unload at maximum speed. Parallel to the dock a conveyor belt transports the concentrate from the weighing hoppers to the concentrate storage building.

ELECTROLYTIC ZINC PLANT OF RUHR-ZINK GMBH 253

The dock i s also equipped for loading sulphuric acid into barges and loading o r unloading other materials such a s leach residues, metals and other bulk products. It must be mentioned that all the leach residue leaves Datteln by barges. The entire loading and un- loading operation i s handled by one operator.

The large concentrate storage building was designed to hold 6 5,000 tons of raw material. The large holding capacity was ne- cessary to handle interrupted shipments from Canada, which are caused by the freezing of the St. Lawrence waterway during winter months.

From the belt conveyor system which i s supported by the roof structure of the storage building, the concentrate i s unloaded at the desired site by a belt tripper. The material can be stored up to a height of 1 5 meters. Should the concentrate arrive with a moisture of l e s s than 6 percent a sprinkler system adds water to prevent

. dusting.

Each type of concentrate i s stored separately, no partition walls a re necessary. The moisture i s such that an angle of repose of 60 degrees enables the piled concentrates to be reasonably divided.

An overhead bridge crane transfers concentrates from stock and discharges the material onto a belt conveyor. The belt i s located along the side of the building and therefore movement of the crane

. i s reduced to a minimum. In case of a crane failure, a portable in- clined belt conveyor which can be fed by a front end loader i s used and proved to be valuable for an uninterrupted roaster operation.

Roasting and Sulphuric Acid

The roasting plant i s equipped with two Lurgi turbulent layer roasters each having a grate area of 55 square meters, with a load factor of 7 metric tons per square meter of grate area, the two roasters have a designed capacity of 770 metric tons of concentrate per operating day.

The concentrates (shown by Table 1) after beeing screened and the oversize denodulized in a hammermill, i s conveyed by a belt system to the roaster feed bins, one each per roaster, having a hold- ing capacity of approximately 100 metric tons each. A pan conveyor with a variable speed and depth regulator transports the concentrate at a predertermined constant rate to the roaster feeding system. The feeding system consists of a disc feeder to equalize the effect of the surges caused by the pan conveyor and a pair of slinger belts discharg- ing directly into each roasting chamber. Usually the two slinger belts

for each roaster a r e in service in o rder to assure an even ternpera- ture distribution throughout the bed.

The temperature in the roas te r i s controlled by the quantity of a i r blown through the grate plate, by the cooling coils in the roas te r bed and if necessary by water injection. Cooling coils were not provided with the original roasting equipment but were added at a la ter date.

The cooling coils a r e connected to the waste heat boiler system.

The particular behaviour of the Canadian Timmins concpntrates (shown by Table 1) require a very close temperature control in the roas te r bed within a range of 20 degrees Centigrade, between 900 and 920, otherwise the bed has a tendency to form agglomerates and thereby lose uniform turbulence.

Bed fluidity i s a lso influenced by the moisture content of the concentrates fed. The ideal moisture fo r Canadian Timmins concen- t ra tes i s about 6 percent. At this moisture only a minor portion of feed enters the bed. About 90 percent o r more will be carr ied by the a i r to the upper portion of the roasting chamber where combustion continues to take place. With 6 percent moisture the bed temperature i s approximately 910 degrees Centigrade and the outlet temperature of the roas te r will show 960 to 980 degrees C e n t i g r a d e . ~ ~ ~ concentrates, with higher than 6 percent moisture these temperature

differences will disappear o r be reversed.

As most of the concentrates delivered have a moisture of 8 percent o r more, additional cooling of the bed had to be provided. Each roaster therefore was equipped with cooling coils, removing 4 million kilo- gram calor ies per hour. At normal throughput ra tes of 350 tons per day per roaster , moisture in the range of 9 percent can be tolerated.

The method of cooling the bed with excess a i r was not successful and increased the already excessive sulphate content of the calcine t o still higher values. At a minimum a i r flow required for proper turbulence in the bed of 28, 000 cubic me t e r s per hour, the total sulphur was 2. 5 percent of which 2.4 percent was sulphate. These results were obtained with a gas composition of 12 percent sulphur dioxide and 3. 5 percent oxygen at the roas te r outlet. Recirculation of calcine was also t r ied with good results for bed cooling but increas- ed the sulphate sulphur content of the final calcine.

The roaster gases after leaving the furnace a r e cooled in a waste heat Lamont type boiler. As excess steam i s sold to a power company .the steam produced was superheated to 400 degrees Centigrade at a p ressure of 50 atmospheres t o meet power plant specifications.

Thus the boiler i s divided into evaporator and superheater sections. F o r each roaster an a rea of 441 square meters for evaporator coils, and 104 square meters for superheater coils a r e provided. Passing through the boiler the gases a r e cooled down to 350 degrees Centigrade. The boilers have a mechanical cleaning mechanism but due to the nature of the dust, hand blowing particularly of the first section of the boiler has to be done once every 24 hours. Low moisture in the concentrate and a minimum amount of water injected into the roaster gives a minimum of accretions and buildup in the waste heat boiler.

After the boiler the gases pass through cyclones and an inter- mediate fan. The gases discharging from the cyclones a r e divided and enter 3 hot cottrells operating in parallel to give maximum flexibility.

With the intermediate fan the draft in the roaster can be regu- lated and what i s considered essential, the hot cottrells can be operated at a slight pressure of 10 millimeters of water. TO this fact the trouble f ree operation of the hot cottrells can be attributed. No accretions o r build up of dust on electrodes have been observed to date.

After the hot cottrells the gases have a dust load of 60 milligrams per cubic meter and a temperature of 330 degrees Centigrade. Six venturi scrubbers a re used to cool and humidify the gas. The final cooling to 38 degrees Centigrade i s done in six horizontal and six vertical lead heat exchangers. The remaining dust together with mist i s removed in a two stage wet cottrell system.

The sulphur dioxide content of the gas i s diluted to 8 percent with air . The gases a r e treated in a Lurgi-Bayer double catalysis acid plant with a conversion efficiency of 99. 5 percent. This single unit has a rated capacity of 645 tons of 100 percent acid per day and the most outstanding feature i s i ts flexibility. During the s tar t up period it was operated for a period at 33 percent of i ts rated capacity without using the preheater. Since the plant has reached i ts full zinc production, the acid plant i s working often at higher than designed throughput without a notable decrease of conversion efficiency.

The acid i s stored in three storage tanks of 15,000 tons holding capacity each. From there it can be pumped to either railroad, highway o r water loading points. As the bulk of the acid leaves Datteln by barge, the loading facilities a r e equipped to load at the rate of 200 tons per hour.

The calcine (shown by Tables 1 and 2) collected from each roaster, waste heat boilers and cyclones i s cooled in an indirect rotary water cooled, heat exchanger, ground in a ball mill and transported by a redler conveyor to a small surge bin. The cottrell dust i s added after the calcine i s discharged from the ball mill. A s ta r feeder regulates the flow of calcine from the surge bin to a Fuller-Kenyon pump. With an a i r pressure of 1.8 atmospheres this pump delivers the fine ground calcine at a rate of 28 tons per hour to either the storage bins or directly into the feed bins in the leaching section.

A small quantity of condensate from the venturi scrubbers and the gas coolers i s withdrawn from the wash circuit, neutralized with soda ash and discarded into the sewage system. Usually this acidified water contains only 5 grams per l i ter acid, 3 - 5 grams per liter chlorine, 1 - 2 grams per l i ter fluorine and 20 milligrams zinc. A recovery system for zinc would not be economical a s the total volume per day i s approximately 10 cubic meters.

The cooling water for the roasters and acid plant i s withdrawn from the adjacent canal. After passing a mechanical screen it i s used without further treatment.

The cooling in the acid plant i s done in a closed circuit and only about 200 cubic meters per hour make up i s continuously added. The cooling water i s circulated over forced draft atmospheric cooling towers.

Leaching Section

The leaching building measures 70 meters long by 45 meters wide. The lower section of the building up to a height of 4. 3 meters i s of concrete. Above this level the building i s constructed of steel. The steel section of the building i s covered and roofed with Robertson Galbestos. The ground floor i s finished with concrete and suitably sloped to drain to a central sump. All spillage, etc. , i s recove red and returned to the solution circuit with pumping equipment. All other operating buildings have similar spillage recovery systems.

The leaching i s conducted batchwise in five mechanically agitated tanks. These leaching tanks have a diameter of 7.60 meters and a height of 3. 20 meters, giving an operating capacity of 130 cubic meters. The self supporting tanks are made of stainless steel No 443 5 according to DIN standards. The thiclmess of the bottom plates a re 6 millimeters, and the walls 5 millimeters. Particularly for the jarosite leaching a thermal insulation of 80 millimeters thickness fibre glass mats a re provided for the side walls. The

ELECTROI-YTIC ZINC PLANT OF RUHR-ZINK GMBH

T A B L E 2

Typical Sc- Analy#is of Calcine

T A B L E 3

Impure Solution Neutral Solution Spent Elektmlyts

N i mg/l

Aa mg/l

C l mg/l

M a dl

CaO a/l

dl

2 - 3

0.6

50 - 100

N.A.

N.A.

N.A.

0.05 - 0.15 < 0.02

50 - loo

4.3

0.7

2.0

< 0.05

< 0.02 I

50 - 100

4.8

0.5

2.0

tank tops a r e covered with plywood segments, supported by the agitator bridge. The lightning agitator i s a propellor type, the blades have a diameter of 1. 780 m e t e r s and rotate at approximately 45 revolution p e r minute and i s driven by a 2 2 kilowatt motor. Four steam jets using 5 atmospheres p ressure and 180 degrees Centigrade steam have rapid heating capacity. A polyester fibre glass stack with a diameter of 600 mil l imeters provides adequate ventilation.

Calcine i s fed f rom two charge bins which a r e suspended on load cells. A predetermined quantity of calcine can be accurately fed by automatic controls. A screw conveyor system with variable speed feeds up t o 100 tons p e r hour of calcine to each one of the leach tanks. The manganese dioxide used fo r oxidation of iron and sulphide sulphur i s transported from a 50 tons holding bin using a s imilar conveying system to each of the leach tanks.

At present (January ' 70) two leaching procedures a r e in use. One system uses 55 percent of a l l calcine and i s leached according to a standard practice, known a s a semi reverse leach. The remain- ing 45 percent a r e leached using hot strong acid and the i ron i s p r e - cipitated a s jarosite. This technique developed by Ruhr-Zink i s based on the patents of Norzink, EZ Industries and Companie Royale Asturienne de s Mines.

F o r the standard leach about 30 percent of the operating tank volume i s filled with wash water and spent electrolyte (shown by Table 3), depending on the sulphide sulphur content of the calcine from 50 to 100 kilos of manganese dioxide (74 - 76 percent manganese dioxide) a r e added. With live s team the temperature of the solution i s brought up to 50 degrees Centigrade. A predetermin ed quantity of calcine i s fed to reduce the f ree acidity to maximum of 15 g rams p e r l i ter . Subsequently spent electrolyte and calcine a r e added simutaneously until the leach tank i s filled. The f ree acid should never exceed 15 g rams per l i ter . The solution i s allowed to react fo r one hour, whereby the acidity will decrease to approximately 5 g rams p e r l i ter . With smal l quantities of calcine additions the pH of the s l u r r y i s brought to 3. 5 - 4. 0. The time required for one leach cycle including pumping t ime i s 2. 5 hours.

During the leaching period the temperature i s maintained at 96 - 98 degrees Centigrade. This high temperature combined with the careful control of the acidity during the entire leach cycle i s essential f o r a zinc extraction of 91 - 91. 5 percent (based on zero water soluble zinc in the residue). Fur thermore the sedimentation rate in the leach pulp clarification system i s such that a minimum amount

ELECTROLYTIC ZINC PLANT OF RUHR-ZINK GMBH 259

of solids a re contained in the thickener overflow. Further filtration prior to purification i s not required. The iron precipitated i s present in the leach residue in the form of iron hydroxide and also in appreci- able quantities a s basic iron sulphate. The specific gravity of the leach solution i s 1. 5 kilograms per liter, corresponding to a zinc concentration of 180 grams per liter. This high zinc concentration i s beneficial for the leach plant capacity. One batch leach dissolves 15 - 16 tons of zinc.

Like many other electrolytic zinc plants, during the start up period, Datteln was faced with an unbalanced sulphate condition. Sulphate in- creased in the electrolyte system in greater quantities than was re- moved by the leach residue. The lmown method of neutralizing excess electrolyte in the leach tanks with lime rock caused considerable troubles in the leach residue handling system. Through sales of zinc sulphate solutions to the Lithopone industry, the sulphate equi- librium was controlled.

In January 1967 some details of a process being developed at the Det Noske Zink Company' s Odda plant became available. From this information Ruhr-Zink adopted a procedure which apart from increasing the leach extraction of zinc, cadmium and copper, also solved the sulphate problem.

The new leaching practice i s conducted by charging a leach tank with electrolyte and calcine added simultaneously until the tank i s two thirds full, resulting in an acidity of 15 grams per liter. The reaction i s carried on for one hour. The free acid i s then increased from 80 to 100 grams per l i ter by the addition of spent electrolyte. After an agitation time of 2. 5 hours the acid concentration i s reduced to about 30 grams per liter. With predetermined small quantities of calcine the acid i s brought slowly to 10 - 20 grams per l i ter and the required sodium ions for the jarosite formation a r e added as a dilute sodium hydroxide solution to the leach tank. During the precipi- tation of the jarosite the free acidity i s slowly reduced by further additions of small lots of calcine. Within 2 hours the iron in solution decreases to 1 gram per l i ter o r less and with additional calcine for neutralization the reaction i s terminated at a pH of 3. 5 - 4.0. The temperature i s maintained between 95 and 98 degrees Centigrade fo r the entire cycle. Zinc leach extraction i s 94. 5 percent and sedi- mentation rates a re slightly inferior compared with a standard neutral leach.

This new leaching procedure solved effectively the unsatisfactory sulphate balance. However, in spite of the better zinc extraction, the economical benefit i s relatively small. An additional step i s

planned for the immediate future. Before the jarosite i s precipitated, the solids in the pulp will be separated in a thickener thereby recover- ing the bulk of the lead, tin and precious metals present in the calcine in a residue not contaminated with jarosite precipitate. This further improvement i s not expected to increase the zinc recovery but the revenues from the sales of the acid insoluble residue will be more substantial.

The manganese concentrations in the plant solution are controlled, by occasionally conducting manganese leaches using manganese dioxide zinc concentrate and spent electrolyte at temperatures up to 95 degrees Centigrade. About 65 cubic meters of spent electrolyte are agitated with 9 tons of manganese dioxide and the corresponding amount of concentrates. After a reaction time of 3 - 4 hours 95 percent of the manganese is dissolved. The reaction can be followed by the decrease of the free acid. After having completed the desired reaction, the leach i s continued by filling the free space in the tank with spent electrolyte and adding calcine to adjust the pH to 3. 5 - 4.0.

Leach Residue Se~aration-Filtration

The hot leach pulp i s pumped from the leach tanks through a surge tank and into the clarification thickener. The surge tank having the same dimensions and equipment a s the leach tanks is needed to regulate the flow of leach pulp to the thickener. When the pulp reaches the thickener, a further neutralization has taken place, the pH i s now 4.0 - 4. 5. The clarification thickener has a diameter of 21.3 5 meters. By gravity the clear overflow (shown by Tables 3 and 4) is fed to the heating tanks of the purification section while the thickened underflow is washed by counter current flow through two additional thickeners each having a diameter of 15. 25 meters. A

8L total of 15 grams per cubic meter floculating agent Sedipur T F , a product of the Badische Anilin und Sodafabrik, is used. The thickener tanks a re made of wood, all submerged parts of the thickener mechanism are of stainless steel. A steel superstructure spans the top of the tank, supporting the mechanism. For reducing heat losses the surface of a l i thickeners a re covered by floating styrofoam sheets.

The underflow of the second washing thickener i s pumped to vacuum drum filters. A total of 3 units a r e installed, each having a filter area of 50 square meters. The filter rates a re 150 - 200 kilograms per

square meter per hour. Essentially all evaporation losses of the entire process a r e replaced by washing the residue on the drum filters with water. Condensate from the heating coils in the purification section and preheated city water a re used for this purpose.

ELECTROLYTIC ZINC PLAYT OF RUHR.ZINK GMBH

T A B L E 4

Thickener I Thickener I1 Thickener 111 DIlLn Filter

Solids g/l

Depth of Clear cm +- - p-

mm. O C

sp. Or. d l t-- I UNDgRKOW

Solids g/l

sp. Or. d l +

2.2

26

45

1.070

850 - 900

- 1.850

1.2

45 - 50

I 4 0.1

150

40

45

1. OgO

300 I - 1.450

- -

< 0.1

1%

70

50

1.200

)oo

- 1.550

- -

< 0.1

150

160

75

1.370

)oo

- 1.650

- -

( Under normal conditions approximately 4 tons of water for each ton of zinc produced are required to compensate for evaporation losses including electrolyte cooling. Spray washing on the drum filters re - duces the water soluble zinc in the leach residue to 1. 2 percent o r less . Filtrate and wash water from the drum filters a r e pumped into 'the second thickener from where they flow counter current to the under- flows from the first and clarification thickeners.

The vacuum drum filters a re made of stainless steel. All piping i s external, facilitating cleaning and maintenance. The polyproplyene fil ter cloth is treated on one side with silicones, facilitating the dis- charge of the cake. Filter cloth life so far averages 3 months.

1 The cake from the f i l ters i s transported by a belt conveyor to a repulping tank. By rigorous agitation, the plastic residue i s repulped and t ransferred by a rubber lined centrifugal pump to the leach residue drier. The rotary drum drier has a length of 18 meters and a diameter of 2 meters. In the first part of the drum chains a r e used to prevent residue adhering to the walls.

The original dr ie r was provided with internal braces to facilitate drying. Later these braces were removed because they interfered with the passage of residue through the drum. The shell is of mild steel having a wall thickness of 12 millimeters. An oil burner consum- ing 500 kilograms of fuel oil per hour provides the necessary heat t o evaporate 6 tons per hour of water. The exhaust gases a r e cleaned in high efficiency cyclones.

Purification

The clear overflow (shown by Table 3) from the clarification thickener i s preheated in three storage tanks. These tanks a s well a s the purification tanks a re constructed of special selected Oregon pine. The diameter of the tanks i s 10 meters and the height 3.1 meters. The active available volume is 170 cubic meters. All tanks a re covered with a plywood covers and ventilated through a 800 millimeters diameter stack. A positive draft is maintained by a dual set of a i r injectors. The agitation is supplied by a similar propeller type agitator as used in the leach tanks. The preheating tanks have steam coils made of stainless steel, in the purification coils of copper a re satisfactory.

The preheated solution is purified in two stages. In the first operation copper, cobalt and nickel a r e removed. In the second stage cadmium and thallium a re eliminated from the solution. These operations a r e again conducted batchwise. This method proved beneficial fo r Datteln, particularly a s cobalt in the impure solution

ELECTROLYTIC ZINC PLANT OF RUHR-ZINK GMBH 263

may vary between 5 and 25 milligrams per liter.

Air atomized zinc dust, copper sulphate and crude arsenic trioxide are used a s reagents. In the first stage zinc dust with a grain fraction of 0.15 - 0. 3 millimeter i s used; for the cadmium removal stage finer zinc dust passing through a 0.15 millimeter screen i s employed.

During the filling of the f i rs t stage purification tank 10 kilograms of arsenic trioxide and 200 kilograms of coarse zinc dust a re added to the agitated solution. When the tank i s filled another 100 kilograms of zinc dust and 25 - 75 kilograms of copper sulphate a re fed. The pH i s controlled at 4. 0 by continuous additions of clarified spent electrolyte. After one hour's reaction time the first determination for cobalt and arsenic i s made. Usually another addition of 100 kilograms of zinc dust i s required to bring the cobalt to the desired concentration of less than 0.2 milligram per liter. At this stage the arsenic i s entirely removed and after a safety addition of another 25 kilograms of zinc dust the so1ution.i~ pumped to the filter presses. The entire first stage purification requires 2 hours. The temperature i s maintained at 95 degrees Centigrade in order to prevent cadmium from coprecipitating.

A total of 5 filter presses a re provided for the separation of the solids. One filter press i s provided a s a spare. Each press has an effective filter a rea of 43 square meters. The material of construction i s bronze. ~ o l y p r o ~ ~ l e n e felt-type filter cloth i s used which i s covered with kraft paper. Filtration proceeds at a rate of 700 to 900 li ter per minute per press at a pressure of 5 atmospheres. Usually a filtration cycle requires 50 minutes using 4 presses in parallel.

For the second stage purification the temperature i s lowered to 7 5 degrees Centigrade in order to precipitate the cadmium. In the tanks of this stage water cooling coils are installed for temperature reduction. This water i s subsequently used for leach residue wash- ing. When the desired temperature i s reached the solution i s acidified with spent electrolyte to a pH of 3. 2 - 3.4 and a first addition of 100 kilograms zinc dust made. If high purity zinc dust i s used, a small addition of copper sulphate acts a s an activator. After half an hour another shot of 100 - 150 kilograms zinc dust i s added. The f i rs t test for cadmium i s made after one hour reaction time. When the solution i s found to contain less than 0. 2 milligram per l i ter cadmium a final safety addition of 50 kilograms of zinc dust i s made and the solution pumped through filter presses into the check tanks. Filter presses for second stage have the same number and dimensions a s for the first stage, they a re however made of an

aluminium-silicon alloy. In the check tank a final analytical de- termination is done for cobalt, arsenic and cadmium by the operator before pumping the solution (shown by Table 3) to the neutral storage tanks. The check tanks have the same dimensions a s the purification tanks while the neutral storage can accommodate 1.100 cubic meters of solution. The second stage purification usually takes 1 1/2 hours.

The purification residue produced during the first stage contains 3 5 - 40 percent copper, 1 percent cobalt, 5 percent arsenic. It is sold to a copper smelter. The residue from the second stage with 15 - 22 percent cadmium and 55 percent zinc is treated for cadmium recovery.

Cadmium Recovery

The fresh second stage purification cake i s first dissolved with spent plant electrolyte in a 50 cubic meters agitator tank. When essentially all cadmium has been dissolved, the solution i s filtered and i ts cadmium content precipitated with the stochiometric quantity of zinc dust. The resulting cadmiurnsponge i s washed, dried and dissolved in spent cadmium electrolyte. The neutral cadmium solution i s subjected to a thallium and copper purification before electro- lyzing. A total of 12 cells, identical to those used in the zinc tank house and operating at current densities up to 100 amperes per square meter, a r e fed with an electrolyte containing 60 - 80 grams per l i ter cadmium, 40 grams per l i ter zinc and 170 grams per l i ter acid. The cathodes a re stripped every 16 hours, washed, dried and sold a s such o r melted under a cover of caustic soda and cast into shapes.

Electrolysis

The tank house is divided into three electrical units. Each electrical circuit consisting of 168 cells with a silicon rectofier unit rated at 30, 000 amperes and 6 50 volts. The pr imary voltage from the utility company of 35,000 volts i s directly connected to the transformers.

The concrete, lead lined cells measure 0.8 meter inside width, 3. 338 meters length and 1.45 meters depth. They a re placed side by side in rows of 14. Every 35 days one double row i s taken out of service and cleaned in from 6 to 10 hours. Each cell accommodates 40 cathodes and 41 cast anodes of corroding grade lead alloyed with 0. 75 percent Ag. The anodes measure 965 millimeters x 550 milli- meters x 8 millimeters while the aluminium cathodes measure 1, 122 millimeters x 600 millimeters x 5 millimeters. The anodes

ELECTROLYTIC ZINC PLANT OF RUHR-ZINK GMBH 265

and cathodes r e st on porcelain insulators. These special shaped insulators act a s spacers a s well a s supports for the electrodes. On the bottom of the anode two porcelain buttons keep the cathode properly spaced.

Electrolyte circulation for each cell i s continuous at a rate of 60 l i ters per minute. The effluent from the cells i s collected in individual launders and flows by gravity through the main collect- ing launder to the spent electrolyte storage tanks. From these tanks the solution i s pumped over an atmospheric cooling tower and flows by gravity into additional electrolyte storage tanks. The electrolyte i s returned to the tank house through an open launder feeding the 36 rows of cells by the way of individual feed launders.

The electrolyte cooling system has a total of 4 wooden tanks, each having a holding capacity of 550 cubic meters. The wooden tanks a re lined with 5 percent antimonial lead. The entire launder system i s of glass fibre reinforced polyester. All pipes used in this section are of polyethylene.

The spent electrolyte (shown by Table 3) i s cooled in a forced draft atmospheric cooling tower which does not contain packing. The tower i s divided into 10 individual units which can be operated independ- ently. The wooden structure i s lined with polyethylene sheets and the bottom pan i s lead lined. The a i r velocity through the units i s 3 meters per second. Three rows of mist eliminators prevent any acid carry- over into the atmosphere. The gypsum precipitated in the tower i s cleaned every 2 weeks, the sludge in the electrolyte storage tanks every 3 months.

The addition of neutral solution to the spent electrolyte i s pumped continuously into one of the electrolyte storage tanks. Reagents re - quired to minimize acid mist formation in the tank house and to de- press lead in the cathodes a re fed continuously into the main elec- trolyte launder. Spent electrolyte i s withdrawn from the electro- lyte storage tanks and pumped to the leach section a s required.

The ampere load to the tank house i s varied throughout the 24 hour period. During the off peak nighttime periods reduced power cost rates a re available. Current densities at the cathode surface vary from 500 to 600 amperes per square meter for the daytime periods. Current densities a re increased to 750 amperes per square meter o r higher during off peak periods which also include week ends. Cell temperatures of the spent electrolyte fluctuate between 39 to 40 degrees Centigrade. However, in the summer of 1969 temperatures of 45 degrees Centigrade were experienced without any detrimental effects on the current efficiency o r cathode

purity. Normal results a re shown by Table 5.

Deposition time i s 24 hours, except for the rows being cleaned, in which the cathodes remain from 38 to 40 hours in the cells. The stripping of the zinc deposit is done manually. An overhead a i r hoist removes 20 cathodes at one time from one cell. The load i s moved, using a monorail to a stripping rack to the main floor of the tank house. The zinc i s pulled from the aluminium cathode by a hand tool and stacked in racks. Two fork lift trucks transport the stripped zinc to the cathode storage a rea after the weight is recorded on a print- ing type scale.

The stripping of the cathode zinc i s completed in 8 hours o r less. One stripper completes the removal of zinc from one row of 14 cells. The weight varies between 9 and 10 metric tons. A stripper i s per- mitted to leave the plant a s soon as his allotted cells a re stripped. A premium is paid for increased current efficiencies based upon the monthly average of his allotted row. Premiums may increase the basic pay as much as 30 percent.

Melting and Casting

After storing the cathodes for 48 hours, they a re melted in one 1800 kilowatt electric induction furnace. The melting rate i s 18 tons per hour. Normal results of the melting and casting department a r e shown by Table 6. The inductors a r e cooled with air. The cooling blowers a r e located in a closed room in order to lower the noise level at the melting furnace and to supply clean a i r to the in- ductors. The exhaust a i r i s kept at 50 degrees Centigrade to pre- vent moisture deposition on the inductor coils.

Dross i s normally skimmed from the surface of the molten metal every 2 to 4 hours using ammonium chloride pellets a s fluxing reagent. The hot dross i s treated in a rotary liquating furnace. The recovered metal is used for zinc dust production and the oxidized dross i s sold. The gases from the liquating drum are filtered in a bag house which i s not satisfactory and a venturi scrubber will be substituted.

Only one quality of zinc i s cast on a straight line, inclined casting machine, having 160 molds. An a i r driven centrifugal pump made of graphite delivers the molten zinc to the casting machine. A variable speed a i r drive for the pump and the velocity of the casting machine can be adjusted to produce a uniform 25 kilograms slab. These slabs a re stacked by hand, cooled for 12 hours weighed and transferred to the metal storage building. For continuous galvanizing , specification metal i s cast in 1 ton blocks. Die casting alloy, containing 4 percent

ELECTROLYTIC ZINC PLANT OF RUHR-ZINK GMBH

T A B L E 5

E l e c t r o l y s i s Data October 1969

Cethode Zinc produced

Loss t ime f o r c lean ing t h e c e l l s

Curren t e f f i c i e n c y

Average cu r r en t Load p e r u n i t (168 c e l l s )

Average Voltage a t r e c t i f i e r

Average Kilowatt hours pe r t o n of cathode

Average cu r r en t dens i t y

Average spent e l e c t r o l y t e temperature

Average e l e c t r o l y t e temperature

Cum n rab i c used per t o n of cathode

Sodium s i l i c a t e used per t o n of cathode

Cre sy l i c a c i d used p e r t o n of cathode

Strontium carbonate used p e r t o n of cathode

Resul t s of t h e Melting and Cast ing Department October 1969

Cathodes melted

kW Hr. per ton of cathode melted

Zinc produced per ton of cathodes m l t e d

Dross

Required Zinc dust

Anwniumehlsride used per t on of cathodes melted

Slab Zinc spec i a l high-grade

Zamsk (d ie c a s t i ng a l l o y )

9.704 t o n s

12.3 h

91.3 S 23.457 A

576 v 3.239 kwh

597 ~ / m ~

39.4OC

34.2Oc

0.031 kg

0.9 kg

0.021 kg

5.74 kg

10.060 tons

103

250 tons

514 tons

1.05 kg

7.374 tons

2.074 tons

aluminium and 1 percent copper i s produced in two 300 kilowatt induction furnaces, holding each 15 tons of metal. The molten zinc i s pumped from the cathode melting furnace into the die cast unit with an asbestos cloth lined steel launder. Load cells on the die cast furnaces a r e used to determine the weight. A master alloy of 80 percent aluminium and 20 percent copper is melted in two oil fired retorts and poured directly into the die cast furnaces. After agitating for 10 minutes the finished alloy is cast on a straight line casting machine at a rate of 6 tons per hour.

Other special die casting alloys a re produced in small quantities. They are prepared in melting kettles and cast by hand into a variety of different shapes.

Modifications

Since the start-up of Datteln plant the excess of sulphates was a serious problem. In the roaster itself an improvement i s obtained, i f the sulphur dioxide content of the roaster gases can be further increased. At present the limiting factor i s the sulphuric acid plant capacity. It is planned to add another unit of 265 tons per day capacity.

In the leaching section the jarosite process will also help to establish a sulphate equilibrium. Once this improvement i s in operation, the possibility of transforming the batch process into a continuous process will be studied. No economical gains can be expected but the work load of the operators could be eased.

TABLE 7

The total employment i s 414 persons and i s divided up a s follows.

Roasting & Sulphuric acid Labour Staff - - Total

3 7 6 43

Leaching & Purification 44 7 5 1

Cadmium Recovery 15 1 16

Tank House

Melting & Casting

General Services

Maintenance 4 2 8 50

Administration