UBC Materials EngineeringMTRL 467Final Design Project Report

Processing of Zinc Plant Residue

Dishoo Randhawa

Estelle Marjorie Nathan

Jeff Jin

Muhammad Harith Mohd Fauzi

Nadeeshika Wickrama Arachchi

Taehyun Yoon

Dr. Edouard Asselin - Group 1

April 7th, 2015

EMAIL

309-6350 Stores Road, Vancouver, BC V6T 1Z4

Design Report

Table of Contents

1.0 EXECUTIVE SUMMARY 3

2.0 PROBLEM SPECIFICATION 5

3.0 TECHNICAL REVIEW 6

4.0 DESIGN OPTIONS 11

5.0 DETAILED DESIGN 13

5.1 MASS BALANCE AND HEAT BALANCE 135.1.1 WATER WASHING 135.1.2 AMMONIA LEACHING 145.1.3 COPPER SOLVENT EXTRACTION 155.1.4 COPPER ELECTROWINNING 165.1.5 CADMIUM RECOVERY 175.1.6 ZINC PRECIPITATION 185.2 MAJOR EQUIPMENT SIZING 195.2.1 FEED FLOW RATES AND TANK VOLUMES 205.2.2 ELECTROWINNING CIRCUIT 22

6.0 SOCIO-ECONOMIC ASSESSMENT OF DESIGN 25

6.1 ECONOMIC ANALYSIS 256.2 ENVIRONMENTAL IMPACT 28

7.0 RECOMMENDATIONS 30

8.0 PROJECT PLANNING 31

9.0 REFERENCES 33

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1.0 Executive Summary

Introduction

Accuracy in developing hydrometallurgical processes are crucial for efficient metal production. Processing a plant can generate highly detrimental issues, if not operated correctly. As a team representing a consulting engineering company, we will assist a client who requires actionable information on whether if it is economically feasible to build and operate a plant in Richmond that processes Cu cement.

Limitation

Hudbay, a major zinc producer, has agreed to sell our client 2000 tonnes per annum of Cu cement cake at a yet to be determined price. Additionally, Hudbay is also willing to sell our client 1200 tonnes per annum of cobalt removal cake. The Cu cement cake consists of 75% metallic Cu and the Co cake contains up to 40% metallic lead, 10% cobalt and appreciable amounts of copper. These cakes contain significant amounts of cadmium, which is a pollutant and must be handled and disposed of per protocol. A new processing method was later identified as a model that maintains low processing costs, wastes less precious metals, and increases revenue as profitable metals were being recovered.

Problem Specification

Many recoverable metals are rejected in the by-product residue, during zinc hydrometallurgical processing. This waste usually contains Cu, Co, Zn, Pb and Cd. If these impurities are present in the electrowinning stage, they would cause detrimental operational issues. Cadmium and lead can cause major environmental concern if not detoxified properly. The task of this project is to propose a model and determine the maximum purchase price that would be affordable to process these residues, given the composition minerals for copper cement of 65.3% copper metallic, 6% zinc sulphate solid, 5.7% zinc metallic, 8.5% cadmium sulphate solid, 3.6% lead metallic, and 10.9% water vapour.

Technical Review

Traditional roast-leach-electrowin process was examined to help provide additional information on various methods for detoxifying cadmium-containing residues. Numerous methods have been analyzed for cadmium-containing waste to be handled carefully and disposed-off properly. In our project, alpha-nitroso-beta-naphthol was selected to precipitate cobalt from raffinate. It is reported that once alpha nitroso beta naphthol dissolved is added to the zinc sulfate solution, 99% of the cobalt is precipitated. EMEW technology was studied for the recovery of copper because of the several advantages that EMEW would provide such as lower operating costs, high mass transport capabilities, easy maintenance, and better control over gaseous products.

Design Options

This section provides flowsheets of three different design models. Each model was examined by carrying out preliminary mass and heat balances. Specifying the appropriate reagents for each option is important at the preliminary stage. Pros and cons for each model were later identified. Based on the economic analysis for each design option, the best processing method was selected.

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Detailed Design

The model we have chosen is a modified version of our second proposed design option where the modifications specifically pertain to the leaching, copper solvent-extraction, and zinc recovery steps. A third-party consultant in the industry suggested an ammonia/ammonium carbonate lixivant be used to maximize the amount of copper recovered. To evaluate if a higher recovery of zinc could be made via precipitation instead of solvent-extraction, we opted to precipitate zinc with sodium carbonate to help recover zinc carbonate and potentially sell it.

Results

The overall cost of the HudBay venture is determined from the project’s capital and operating costs. The calculation for the total capital cost was tabulated in a table that can be found under the economic analysis section. Total capital cost was found to be roughly $9.8 million while the operating cost was calculated to be about $5.6 million per annum. Based on the total operating cost for processing 2,000,000 kilograms of cake per year, we would be able to purchase the cakes from HudBay at about $2.79 per kilogram of cake. The revenue obtained per kilogram of cake is $4.11. Each year, we can expect to have a revenue of about $8.2 million. Taking everything into account, the model that we have selected was the best choice where not only copper recovery was maximized to at least 90%, a high revenue can be expected.

Environmental Impact

Processing the copper cement cake will produce a number of items. The copper and zinc carbonate that we are recovering from the cake will generate two waste products which are cadmium and lead. These two products will raise concerns with respect to the environmental impact of our project as they are deleterious substances that must be handled properly so as to not avoid undesirable outcomes. Based on the economic analysis, fees to impound these waste products were calculated for and were included in the overall operating costs of the project.

Recommendations

Examine various ways for the removal of nickel and lead that would be worth recovering. Further design and testing of copper extraction with soluble zinc is recommended. Study different lixiviants that would promote better volume ratio of aqueous to organic phases to help increase in copper stripping efficiency.

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2.0 Problem Specification

As per our client’s (Dr. Edouard Asselin) request our firm (Group 1) has prepared a report regarding the feasibility of a new processing plant in Richmond, B.C. The aim of this new facility would be to process copper cake obtained from HudBay, which operates a zinc pressure leaching facility in Flin Flon, Manitoba. HudBay has agreed to sell our client 2000 tonnes per annum of Cu cement cake at a yet to be determined price. The Cu cement cake contains up to 75% metallic Cu and it contains significant amount of cadmium, which is a pollutant that must be handled and disposed of per protocol.

During zinc hydrometallurgical processing many recoverable metals are rejected in the by-product residue. The residue arises from the cementation process used to refine the zinc solution before electrowinning (Lu et al., 2014). This waste usually contains Cu, Co, Zn, Pb and Cd at varying concentrations depending on the exact refining processes (Lu et al., 2014). If these impurities are allowed to report to zinc electrowinning they cause a lot of detrimental operational issues. Cadmium and lead also present a major environmental concern if not detoxified. It is estimated that for each kiloton of Zinc produced 10 tons of Cd containing residue is also produced (Lu et al., 2014).

The objective of this study is to determine whether it is economically feasible for our client to set up a facility in Richmond to process these products to extract marketable metals and to determine the maximum purchase price that would make it affordable to process these residues. At this stage in time, we are to compare three different design option models’ operating parameters and reagent requirements in order to advise our client on whether to process one or both of the cakes; or not to process at all.

There are three types of design option models that are considered: Cd and Cu recovery, Cd, Cu, and Zn recovery, and Cd, modified Cu, and Zn carbonate recovery. For each, we are tasked with determining the operating parameters such as heat and reagent requirements. Subsequently, with the model that we found the most attractive based on preliminary analysis, we are to provide a process flowsheet, size of equipment and materials, and a capital and operating cost estimate.

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3.0 Technical Review

Overview

The Roast-Leach-Electrowin process is also known as electrolysis process that consists of four steps: leaching, purification, electrolysis, and melting/casting. In the leaching step, the result of the process is a solid that contains precious metals and liquid that contains zinc, cadmium, copper, and cobalt. Then, the liquid leach product needs to be purified in large agitated tanks. The purification process utilizes cementation process to further purify the zinc. Steam and zinc dust is used to remove cadmium, copper, and cobalt from interfering the electrolysis process. The zinc sulfate solution must be very pure for electrowinning to be efficient. If there are impurities, the decomposition voltage will change that will cause the electrolysis cell to produce hydrogen gas rather than zinc metal. Then, copper is extracted by electrowinning.

Ammonia Leaching

From our preliminary analysis, we found ammonia leaching to be the most appealing method to proceed with our project. In 2007, Alexander Mining first developed ammonia leaching known as AmmLeach . The process uses ammonia-based chemistry to selectively extract copper, zinc, nickel, and cobalt from ore deposits. Ammonia leaching has the highest potential to extract valuable metals selectively and cost effectively. The recovery is listed 70-80% with heap lifetime of ~80-130 typical working days as to acid leaching is 80% recovery with heap lifetime of 55-480 days. Additionally, the process is more environmentally friendly than the conventional acid leaching that uses acid to dissolve copper and cobalt. Heap can be washed and left in the tank, and residual ammonia can be used as fertilizer for vegetation regrowth. Further solvent extraction process is simpler and can save time to be followed by a solvent extraction electrowinning process for copper.

Solvent Extraction

Solvent Extraction is one of the unit operations in hydrometallurgical plant where the pregnant leach solution (PLS) is transferred to solvent extraction operation for metal recovery (Davenport, 2002). PLS is initially stored in the PLS pond before being pumped to the loading stage where it is contacted with an organic extractant dissolved in a diluent. Extractant is an aromatic hydroxyl oxime and it is known in the market under the trade name, Lix (Kordosky, 2002). This oxime reagent can be abbreviated as HR where H is an acidic proton. While, the R group is a straight, long hydrocarbon chain and its purpose is to make the extractant hydrophobic and to lower the solubility of the extractant in the water.

The reaction below describes the reaction of loading stage between organic diluent and copper solution:

2RH org + Cu+2aq -> CuR2 org + 2H+

aq

This reaction is between copper sulfate in the form of PLS where it is reacted with the organic extractant forming CuR2, an organic phase of copper and sulfuric acid. This reaction involves metal ions transfer from aqueous phase to organic phase during loading. Since the extractant is very selective, the other metals are rejected in the form of raffinate (Kordosky, 2006). Selectivity is an important parameter is solvent extraction and it is defined as the extent of extraction of one metal ion compared to another. The more selective a reagent to the copper, the less it extracts other metal ions.

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The second stage is called as stripping stage which is just the reverse of the loading reaction:

CuR2 org + 2H+aq -> 2RH org + Cu+2

aq

The goal for stripping operation is to recover as much copper from the organic phase before being transferred to electrowinning (Merigold, 1996).This stage reproduces the protonated reagent and converts organic copper back into aqueous phase. The organic phase is contacted with strong acidic solution known as spent electrolyte coming from electrowinning (EW). High concentration of the spent electrolyte ensures that the stripping curve in the McCabe –Thiele graph to be quite steep for higher concentration of copper in the aqueous phase. It is to ensure a higher degree of metal recovery.

However, there are certain issues to be noticed as aldoximes tend to be strong extractants for copper that they are difficult to strip with conventional lean electrolyte coming from EW (Cotton, 1992). Conversely, the loaded organic would be very easy to strip but that would not be much metal ions in the organic anyway. Therefore, it is essential to use a moderately strong extractant with addition of certain amount of modifier so that we can achieve an optimum level of metal recovery. The degree of extraction is defined as the ratio of what is transferred to aqueous phase over what came in with the aqueous phase as shown in the equation below:

Percentage of extraction=Mass transfered¿ organic ¿Mass enteringthe aqueous

×100

Traditionally, the mixers and settlers are the major equipment used in solvent extraction. In the mixer, the aqueous and organic solutions are contacted to form dispersion so that there would be an increase in the interfacial surface area for faster copper transfer between the phases (“Cognis Group”, 2007). The size of the mixers and flow rates of the solution determine the residence time for each single tank. Typical residence time in the mixer system is about 3 minutes (Davenport, 2002).

After mixing, the phases flow into settler and separate into individual aqueous and organic solution with high extent of purity. Observation shows that the less dense organic stays at the top of settler (Merigold, 1996) while the denser aqueous stays at the bottom of the settler. Residence time in the settler can be calculated by length of the tank divided by the linear velocity of the phase. The diagram below depicts the flow sheet of solvent extraction where copper entering as PLS, being processed and gets ready for electrowinning:

Diagram 1: the arrangement of mixer and settles for copper solvent extraction, Copper Leaching, Solvent Extraction, and Electrowinning Technology,)Lorem Ipsum 8

Mixer

Settler

PLS

Raffinate

Mixer

Settler

Raffinate

Organic, partly loaded

Mixer

Settler

Aqueous, partly

depleted

PLSOrganic,

higher loaded

Organic, highest loaded

Mixer

Settler

Barren organic

Rich electrolyte

Lean electrolyte

P1 E2 E1Stripping

MixerMixer

Settler

PLS

Raffinate

MixerMixer

Settler

Raffinate

Organic, partly loaded

MixerMixer

Settler

Aqueous, partly

depleted

PLSOrganic,

higher loaded

Organic, highest loaded

MixerMixer

Settler

Barren organic

Rich electrolyte

Lean electrolyte

P1 E2 E1Stripping

Design Report

Electrowinning

Electrowinning is a form of electrolysis where metal is plated onto both sides of the cathode sheets while oxygen evolution that oxidizes water occurs at the anode. Enriched electrolyte supplied from solvent extraction (SX) is fed into the cells and returned back to solvent extraction as the lean electrolyte (Prasad, 1992). Once the copper has been plated at the required thickness, the cathodes are removed from the cell and ready for the sale. Electrowinning process is based on Faraday’s Law where the moles of metal produced in electrolysis is directly proportional to the charges passed as described in the equation below:

n × M= qn×F

q = number of charge, F = Faraday constant; 96485 C

Rearranging the equation gives us the amount of mass of the plated metal as shown in the formula below:

M =I∗t∗AWn∗F

AW= atomic weight, t = time, n = number of moles, I = electric current

In addition to that, current efficiency for metal plating then is the ratio of the actual mass of metal plated to the theoretical mass obtained from Faraday’s Law and eventually we obtain the following equation:

CE=100∗n∗FI∗tAW

While the energy efficiency is the ratio of theoretical energy required to the actual energy required in percent and given by the equation below:

EE=(|∆ E|×CE)/∆ Eapplied

The electrodes are connected to an external power supply so that the voltage exceeds and opposes the thermodynamic cell voltage. This phenomenon is also known as over potential.

Let’s look at the reaction for zinc plating as shown below:

Zn2+ + 2e- = Zn

H2 = 2H+ + 2e-

Overall Reactions: Zn2+ + H2 = Zn + 2H+ ∆ E0=−0.76V

This overall reaction is not favorable and requires external voltage to be greater than 0.76 V to force the reaction to be favorable. In addition to that, the rate of metal plating is directly proportional to the current thus it is crucial to have a control over the current efficiency over the electrowinning operation. For copper electrowinning, the cathode copper is plated onto stainless steel about

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3mm thick and 1m × 1m dimensions (Jenkins, 1999). Stainless steel is used to minimize corrosion. Cu2+ as the electrolyte is kept at high concentration to maintain a high current density. While, rich electrolyte coming from solvent extraction has concentration about 25g/L Cu2 and H2SO4 is within the electrolyte about 150g/L.

Copper is usually plated about 5mm for 5 to 10 days and this operation temperature is at 45-50 degree Celsius (Merigold, 1996). Therefore a good control electrolyte temperature is required to obtain a smooth and dense deposit copper.

The rich electrolyte from the solvent extraction is transferred to electrolyte storage tank:

Diagram 2: how the electrowinning operation is connected to the SX stripping for Cu EW, Electrolytic copper -leach, solvent extraction and electrowinning world operating data)

Higher current efficiency, higher current density and higher cathode surface area increase the copper plating rate in the cells (Jenskins, 1990). Acid concentrations are kept high to keep the resistance of the electrolyte lower for greater ionic migration. The electrolyte is heated by steam coils up to 65 degree Celsius entering the circuit and leaves the operation at 60 degree Celsius.

EMEW technology is widely being used for recovery of copper. The technology provides excellent quality copper directly, safely and economically in versatile ways. This technology is capable of recycling of acid, production purity of 99.99% copper cathode as it offers better impurity control. Other advantages are low labor and operating cost because of lesser inventory and lower recycles, lower ventilation requirement, and less maintenance. This technology also delivers continuous and easy control of copper concentration, and reduction in handling and transportation. Additionally, health and safety concerns are neglected as there are no dealings with acid mist, arsine, lead anodes, and more. EMEW technology offers great overall efficiency

Cadmium Removal

The storage of cadmium containing waste awaiting disposal should be under conditions that do not allow it to escape into the environment. When moving the waste around the site care must be taken to avoid spillage during handling and transportation. Any spills or containment breaches that do occur must be dealt with and cleaned up immediately. In addition, preventative measure such as sweeping of the roadways and damping of loose dirt to avoid toxic dust should be performed on a routine basis (OECD, 1995).

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In British Columbia there are strict regulations and standards that define how hazardous waste such as cadmium is transported and disposed of. A Hazardous Waste Transport license and a third-party hazardous waste insurance is required for the transportation of the toxic waste from the plant to the disposal facility (The Province of British Columbia [PBC], 2014). The government also requires the carrier to provide a contingency plan outlining how they will respond to potential emergencies involving accidental release of hazardous waste into the environment (PBC, 2014). After applying for the license one representative from the company must arrange to write the Hazardous Waste Transport License test designed to evaluate the applicants’ knowledge of the hazardous waste regulations (PBC, 2014)

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4.0 Design Options

As a group, we have proposed three different models and performed mass and heat balances for each of the reactions involved. We started off by indicating the composition to be 65.3% of copper, 6% zinc sulphate solid, 5.7% of zinc, 8.5% of cadmium sulphate, 3.6% of lead, and roughly 10.9% of moisture. By writing out the appropriate equations for each of the stages involved in each model, we were able to calculate the mass present for each element based on the mole-to-mole ratio. Knowing the appropriate mass for each species, we moved on to the heat balance portion to determine the total enthalpy for the model.

For design option 1 (cadmium and copper recovery), the cadmium-containing residue will be oxidatively leached with water. Following the water washing stage, the solid residue which contains all the valuable metals such as copper, zinc, cobalt, and cadmium will be leached with sulfuric acid. The remaining solution will then be sent to the copper SX circuit to further separate zinc and cadmium. At this point, copper can be recovered by using an organic substance known as LIX 973. In order to fully eliminate the impurity metals, the leachate must be purified well before electrowinning so as to achieve a high-purity zinc product at high current efficiency (Lu et al., 2014). Based on previous lab results, almost 90% of the cadmium was extracted (Diankun et al., 2014) by undergoing water washing and the extraction of cobalt, zinc, and cadmium had greatly increased when sulfuric acid was used as the lixiviant. However, after much consideration, we decided not to consider this process because the remaining precious metals in the final raffinate go un-recovered and the extraction of copper is relatively low compared to the other three metals when leached with sulfuric acid.

For design option 2 (cadmium, copper, and zinc recovery), it is a continuation of design option 1 that handles the raffinate containing cobalt, zinc, and some traces of cadmium that exit the copper solvent-extraction phase. The first process that occurs after this solvent-extraction process is a precipitation stage during which α-nitroso-β-naphthol is used to precipitate out the cobalt from the raffinate of interest. The other product of this precipitation stage is a solution containing zinc and bits of cadmium. This solution is subjected to another solvent-extraction circuit, this time to recover more zinc. For this zinc solvent-extraction, the lixivant used was P204 (DEPHA). The recovered zinc is then to be sent to electrowinning whereby it will be made it to a more profitable state with which we can increase project revenue. This option maintains low processing costs and the final raffinate does not waste precious metals. Nevertheless, copper is still being recovered in less-than-optimal amounts (58%) and thus, we will not consider this process.

For design option 3 (cadmium, modified copper, zinc carbonate recovery), it is a modified version of design option 2 where the modifications specifically pertain to the leaching, copper solvent-extraction, and zinc recovery steps. A third-party consultant with much experience in the industry recommended that an ammonia/ammonium carbonate lixivant be used to maximize the amount of copper recovered. By preference of our client, we chose the ammonium sulphate; as the copper will be as copper metal and cuprous oxide. Leach recovery is expected to reach over 90%, substantially increasing revenue from the copper from that of which came from a 58% extraction (design options 1 and 2). Since the leach is basic in this model, we chose to use LIX 841 as it is better suited to these conditions as LIX 973 is preferred for acidic leaching. The difference between design option 2 and 3 is the method for zinc recovery. We opted to precipitate the zinc with sodium carbonate to recover zinc carbonate and to sell as it is. This change was made to evaluate if a higher recovery of zinc could be made via precipitation instead of solvent-extraction. It is important to note that there will be some traces of the nickel in the raffinate that we are choosing to deem negligible in quantity and profitability in the scope of our project, and thus we will not consider it. Based on the results that we have obtained for all three different processes, design option 3 seemed to be the best choice since copper recovery is maximized to at least 90% which corresponds to a significant increase in revenue. Having said that, the final raffinate for this process wastes less precious metals and also, it incredibly maintains

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low processing costs. Therefore, as a group, we were all in agreement to proceed with design option 3.

Figure 1. Process flowsheet for Design Option 3

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5.0 Detailed Design

Feed

Species Composition %

Copper metallic 65.3Zinc sulphate 6.0Zinc metallic 5.7Cadmium sulphate 8.5Lead metallic 3.6

Total 89.1Table 1: Feed composition from Hudbay

5.1.1 Water Washing

ZnSO4 (s) + H2O(l) ----> ZnSO4 (aq) + H2O(l)CdSO4 (s) + H2O(l) ----> CdSO4 (aq) + H2O(l)

Diagram 3: Water washing operation

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Input OutputSpecies ZnSO4 H2O ZnSO4 H2OMW 161.44 18.02 161.44 18.02Mass (kg) 6.00 0.67 6.00 0.67

Table 2: the input and output mass for dissolving the zinc sulphate solid

Input OutputSpecies CdSO4 H2O CdSO4 H2OMW 208.47 18.02 208.47 18.02Mass (kg) 8.50 0.73 8.50 0.73

Table 3: the input and output mass for dissolving cadmium sulphate solid

HudBay, a Canadian integrated mining company, had provided this project team with a composition minerals for copper cement comprising 65.3% copper metallic, 6% zinc sulphate solid, 5.7% zinc metallic, 8.5% cadmium sulphate solid and 3.6% lead metallic, as shown in the table above. The final component was water vapour at 10.9%. The team utilized mass and heat balance spreadsheet, provided by Professor Asselin, to solve mass balance involving several processes as shown in the proposed hydrometallurgical flow sheets below. A 100 kg feed basis was used for this mass balance calculation.

Design Report

Total Enthalphies:

Total Mass Input 200 kgTotal Enthalpy -1834 MJTotal Mass Output 200 kgTotal Enthalpy -1834 MJ

5.1.2 Ammonia Leaching

Cu(s) 0.5O2 +(NH4)2SO4 +2NH3 -> Cu(NH3)4SO4 +H2OZn(s) 0.5O2 +(NH4)2SO4 +2NH3 -> Zn(NH3)4SO4 +H2O

Diagram 4: Ammonia leaching operation

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Species Cu (NH4)2SO4 O2 NH3 Cu(NH3)4SO4 H2OMW 63.55 132.14 32.00 17.03 227.73 18.02Mass (kg) 65.30 135.79 16.44 35.00 234.02 18.51

Input Output

Table 4: the input mass and output mass for complexing copper metallic with ammonia and ammonium sulphate

Species Zn (NH4)2SO4 O2 NH3 Zn(NH3)4SO4 H2O

MW 65.38 132.14 32.00 17.03 229.56 18.02

Mass (kg) 5.70 11.52 1.39 2.97 20.01 1.57

Input Output

Table 5: the input and output mass for complexing zinc metallic with ammonia and ammonium sulphate

Total Enthalpies:

Total Mass Input 274 kgTotal Enthalpy -1492 MJTotal Mass Output 274 kgTotal Enthalpy -1680 MJ

Copper concentrates were transferred to a water washing process to dissolve the cadmium sulphate and zinc sulphate solids. However, this process did not dissolve other components as they required stronger dissolving reagents. For mass balance calculation, solubility limit for cadmium sulphate (Ksp) was 76.6 g/100g water while solubility limit for zinc sulphate was 53.8 g/100 g water at 20 degrees Celsius. With these solubility limits, the amount of water to dissolve all sulphate metals was determined to at 0.67kg to dissolve 6 kg zinc sulphate solid and 0.73 kg to dissolve 8.50 kg cadmium sulphate solid. The remaining aqueous solution at this stage was forwarded to the next operation for further processing.

Design Report

The remaining components of the copper cement (the remaining aqueous solution) were transferred to the ammonia leaching process where the 65.3 kg copper metallic and 5.7 kg zinc metallic were reacted with ammonia and ammonium sulphate forming a complex solution. Oxygen was added into the leaching process. This leaching process formed 234.02 kg copper tetra amine sulphate and 20.01 kg zinc tetra amine sulphate. The remaining metals, which were lead and cadmium, did not reacted with ammonia as the solubility of cadmium metallic and lead metallic in liquid ammonia were too low to form amine complexes.

With regard to ammonia and ammonium sulphate, this project team proposed to recycle ammonia produced from zinc precipitation process and electrowinning. The table below shows the ammonia and ammonium sulphate as total input and total output for all processes and recycling ammonia reduces the ammonia consumption on average 59.8 kg per 100 kg feed.

Ammonia

Ammonium Sulfate Total

Input(kg) 52.03 147.3 199.3Output(kg) 9.7 70.02 79.7Required (kg) 42.3 77.3 59.8

Table 6: Ammonia required for ammonia leaching

5.1.3 Copper Solvent Extraction

Diagram 5: Copper Solvent Extraction operation

Loading Stage:

Cu(NH3)4SO4 +2HR -> CuR2 + (NH4)2 SO4 + 2NH3

Input OutputSpecies Cu(NH3)4SO4 HR CuR2 (NH4)2SO4 NH3MW 227.73 115.24 292.01 132.14 17.03Mass (kg) 234.02 236.84 300.07 135.79 35.00

Table 7: copper amine reacting with organic reagent for loading stage

The amine complexes from the leaching process were transferred to the copper solvent extraction process. Pregnant leach solution (PLS) containing copper tetra amine sulphate was reacted with organic reagent, aromatic hydroxime (HR) or in open market was Lorem Ipsum 16

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known as Lix, to form an organic copper. Due to the selectivity of HR, only copper amine reacted while zinc amine was transferred to the next operation (zinc precipitation). The 234.02 kg copper tetra amine sulphate reacted with organic solvent to produce 300.07kg CuR2, (NH4)2SO4 and NH3.

Stripping Stage:

CuR2 +(NH4)2SO4 + 2NH3 -> Cu(NH3)4SO4 +2HR

Input OutputSpecies CuR2 (NH4)2SO4 NH3 Cu(NH3)4SO4 HR

MW 292.01 132.14 17.03 227.73115.2

4Mass (kg) 300.07 135.79 35.00 234.02

236.84

Table 8: organic copper reacting with ammonia and ammonium sulphate to form purified copper amine

Stripping stage is the second stage where CuR2 were reacted with ammonium sulphate and ammonia that were produced from electrowinning process. This project team assumed at this stage that the mass loss for both stages to be zero. Purified copper tetra amine sulphate was then transferred to electrowinning operation. While the raffinate containing zinc tetra amine sulphate was transferred to next process, zinc precipitation.

Total Enthalpies:

Total Mass Input 942 kgTotal Enthalpy -3575 MJTotal Mass Output 942 kgTotal Enthalpy -3575 MJ

5.1.4 Electrowinning

Cu(NH3)4SO4 + H2O = Cu + 0.5O2 + (NH4)2SO4 + 2NH3

Diagram 6: Electrowinning operation

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Copper tetra amine sulphate was one of the electrolyte components, and electrowinning operation aimed to obtain purified metallic copper. The flowrate of copper tetra amine entering the operation is at 41kg/m3 and the flow rate of copper tetra amine sulphate exiting the operation is at 30kg/m3. The change in flow rate enables our project team to calculate the amount of copper plated per operation which is 18kg. Ammonia and ammonium sulphate produced from this operation were recycled to ammonia leaching.

Design Report

Input OutputSpecies Cu(NH3)4SO4 H2O Cu O2 (NH4)2SO4 NH3 Cu(NH3)4SO4MW 227.73 18.02 63.55 32.00 132.14 17.03 227.73Mass (kg) 234.02 5.13 18.09 4.55 37.62 9.70 169.19

Table 9: reactions involved during electrowinning operation with the corresponding mass for each component

Total Enthalpies:

Total Mass Input 239.1896 kgTotal Enthalpy -932.3222104 MJTotal Mass Output 239.1450019 kgTotal Enthalpy -1277.228116 MJ

5.1.5 Cadmium Removal

Diagram 7: Cadmium Removal operation

Total Enthalpies:

Total Mass Input 11.166 kgTotal Enthalpy -38.060158 MJTotal Mass Output 11.167 kgTotal Enthalpy -38.060318 MJ

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Species CdSO4 Zn ZnSO4 CdMW 208.47 65.38 161.47 112.41Mass (kg) 8.50 2.67 6.58 4.58

Input Output

Table 10: the input and output mass for cadmium cementation

Cadmium sulphate was transferred from water washing, after being dissolved to form cadmium sulphate (aqueous) in the water washing operation. The project team proposed to remove cadmium by cementation of cadmium sulphate using zinc dust. 2.67 kg zinc dust was used to remove cadmium from the solution and 6.58 kg zinc sulphate formed from the reaction was transferred to zinc precipitation process.

Design Report

5.1.6 Zinc Precipitation

Zn(NH3)4SO4 + 2Na2CO3 + 2H2SO4 = ZnCO3 + [NH4]2CO3 + 2Na2SO4 + [NH4]2SO4

Diagram 8: Zinc Precipitation operation

Input Output

SpeciesZn(NH3)4SO

4 Na2CO3 H2SO4 ZnCO3 [NH4]2CO3 Na2SO4 [NH4]2SO4MW 229.560 105.988 98.079 125.390 96.090 142.104 132.140

Mass(Kg) 32.59830.10103

527.85484

617.80564

213.6449809

240.3581

318.764156

3Table 11: the input and output mass for zinc precipitation

Total Enthalpies:

Total Mass Input 90.55 kgTotal Enthalpy -773 MJTotal Mass Output 90.55 kgTotal Enthalpy -830 MJ

Raffinate containing zinc tetra amine sulphate was transferred from solvent extraction process, and 32.598 kg of the zinc tetra amine was precipitated by reacting the complex solution with sodium carbonate forming 17.8 kg of zinc carbonate, ammonium carbonate, ammonium sulphate and sodium sulphate. Ammonium carbonate was recycled back to ammonia leaching process. The recycle process was performed by evaporating the carbonate, reacting it with water to form back ammonia and this ammonia was then transferred to ammonia leaching. One issue which the team recommended for future resolution is to find solution for sodium sulphate as it will go through all processes, and may cause problem to the entire operations.

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5.2 Major Equipment Sizing

To size the major equipment of the processing facility we decided that the plant would be running 250 days per year for 12 hours a day. Not operating on weekends and holidays allows us to reduce operating costs, as workers must be paid a premium for working those shifts. The designed capacity of the facility is to process 2000 tonnes of copper cement cake per year, which allows us to process 8 tonnes per day. To achieve this target we determined that it was reasonable to operate the plant only 12 hours a day without sacrificing recovery and allowing for some leeway in the case of unplanned downtime. The proposed equipment diagram is shown below:

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Figure 2. Equipment diagram

5.2.1 Feed Flow Rates and Tank Volumes

The table below shows the calculated flow rate and size of each tank:

Table 12: Tank sizes and feed rates

By utilizing the mass balance, the feed flow rates and volumes were determined for each tank. A few sample calculations for flow rate and tank volume are shown below:

Cement Holding Tank:

Scope: 8000 kg/day of cement cake @ 12 hours per day

Flow rate (mass): 667 kg/hr of cement cake

Cake Density: 3910 kg/m3

Volumetric flow rate: 667 (kg/hr) / 3910 (kg/m3) = 0.17 m3/hr

Tank refilled every: 12 hours

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Tank volume: 0.17 m3/hr * 12 hours / 0.7 = 2.92 m3

* The tank volume is divided by 0.7 to account for a 30% oversize factor for all tanks.

Ammonia Leach Tank:

Solid residue

Flow rate (mass): 710 kg/hr

Density: 3037 kg/m3

Volumetric flow rate: 710 (kg/hr) / 3037 (kg/m3) = 0.23 m3/hr

Ammonia:

Flow rate (mass): 1867 kg/hr

Density: 880 kg/m3

Volumetric flow rate: 2.12 m3/hr

Leaching time: 1 hour

Tank volume: ((0.23 m3/hr +2.12 m3/hr)*1 hr)/0.7 = 3.36 m3

The solid residue flow rate and density are calculated as follows:

Basis: 100 kg copper cement cake

Solid Residue from Water Leach

Cu(s) Zn(s) Pb(s) Water SUM

Mass (kg): 65.3 5.7 3.6 32 107

Density (kg/m3): 3910 1000

Weighted Density: 3037

Water mass (@30% moisture) = [(65.3+5.7+3.6)/0.7]- (65.3+5.7+3.6) = 32 kg

Weighted density = [(65.3+5.7+3.6)/107]*3910 + [32/107]*1000 = 3037 kg/m3

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Flow rate = 107 kg * (667 kg/hr of cement flow rate/ 100 kg basis) = 710 kg/hr

Loading Mixer:

Contact time: 20 minutes = 0.33 hour

A/O: 1

Flow rate: 1.89 m3/hr

Tank volume: 1.89 m3/hr * 0.33 hour / 0.7 = 0.9 m3

Stripping Settler:

Settling time: 1 hr

A/O: 1

Flow rate: 1.89 m3/hr

Tank volume: 1.89 m3/hr * 1 hr / 0.7 = 2.7 m3

All tanks will be constructed of stainless steel to avoid corrosion during the 15-year lifetime of the plant. Ductile iron cased vertical motor centrifugal pumps will be used where necessary as depicted in the equipment diagram.

5.2.2 Electrowinning Circuit

For the recovery of copper the plant will implement Electrometal’s EMEW cells.

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Figure 3. EMEW cells

The advantages of using EMEW technology versus conventional Merrill Crowe tanks are as follows:

1. Lower operating costs (Electrometal Technologies Limited [ETL], 2014).2. High mass transport capabilities – decreased electroplating time (ETL, 2014).3. Modular design – easy maintenance, expansion and relocation (ETL, 2014).4. Closed cell – better control over gaseous products (ETL, 2014).

To determine the number of EMEW cells required the following parameters were used:

Faradays constant (F) 96500C/equivalent

Cathode Area (A) 1 m2

Current Density (J) 400 A/m2

Copper density (D) 8960 kg/m3

Copper Molar Mass (Mw) 0.063546 kg/mol

Current (I = J*A) 400 A

Copper Plating Thickness (T) 0.005 m

Density of Electrolyte 1080 kg/m3

Copper charge (n) 2

Copper recovered (from mass balance): 18.09 kg / 100 kg copper cement cake

Copper cement flow rate: 667 m3/hr

Mass of Copper recovered per hour: 18.09 * (667/100) = 120.6 kg/hr

Mass of Copper recovered per day: 120.6*12 = 1447.2 kg

Mass of Cu plated on one cell (Mc): A*D*T = 44.8 kg

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Cells required for one day’s batch: 1447.2/44.8 = 32 cells/day

Time required to plate one cell: t=Mc∗F∗nI∗Mw

t = 340163 seconds = 3.93 days

Number of cells required: 32 cells/day * 3.93 days = 127 cells

15% Contingency 19 additional cells

TOTAL EMEW Cells Required: 146 cells

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6.0 Socio-Economic Assessment of Design

6.1 Economic AnalysisThe total cost of the HudBay venture is determined from the project’s capital and operating costs. The capital cost is comprised of fixed, one-time expenses such as the purchase of land, buildings, construction, and equipment to be used in the processing plant. Essentially, the capital cost is the total cost needed to bring a project to a commercially operable standpoint. In our case, the capital cost is made in bulk of various direct and indirect costs; to conduct an appropriate estimated cost analysis, each of these costs were assigned to be a certain percentage of the total equipment cost based on recommended values for general solid-fluid processing plants. The calculation of the total capital cost is tabulated in Table 2 below and has been found to be approximately $9.8 million.

TOTAL

$9,804,930.25

I $8,334,190.71i- Low end High end Average Low High Average

1- 100% 100% 100% 1,530,607 1,530,607 1,530,607

2- 25% 39% 32% 382,652 596,937 489,794.08

3- 8% 9% 9% 122,449 137,755 130,102

4- 13% 26% 20% 198,979 397,958 298,468

5- 31% 31% 31% 474,488 474,488 474,488

6- 10% 15% 13% 153,061 229,591 191,325.81

7- 29% 29% 29% 443,876 443,876 443,876

8- Ancillary Building and Services

a. Admins., Shops, Medical, etc 10% 20% 15% 153,061 306,121 229,591

30% 80% 55% 459,182 1,224,485 841,834

9- 4% 8% 6% 61,224 122,449 91,836

10- 0% 0% 0% - - - TOTAL 160% 257% 209% 3,979,577 5,464,265 4,721,921

ii- Low end High end Average Low High Average1- 8% 10% 9% 377,754 472,192 424,972.90 2- 10% 10% 10% 472,192 472,192 472,192 3- 10% 12% 11% 472,192 566,631 519,411 4- 10% 10% 10% 472,192 472,192 472,192 5- 2% 8% 5% 94,438 377,754 236,096 6- 4% 4% 4% 188,877 188,877 188,877 7- 25% 30% 28% 1,180,480 1,416,576 1,298,528.30

TOTAL 69% 84% 77% 3,258,126 3,966,414 3,612,270

II $1,470,739.54Low end High end Average Low High Average

TOTAL 10% 20% 15% 926,021 2,083,548 1,470,740

Indirect Cost

Fixed Capital

Working Capital

TOTAL CAPITAL COST:

Offsite Infrastructure

Land and Site Prep

b. Utilities

Contingency

MODEL 3

Equipment CostDirect Cost

Building and Structure

Electrical Equip. and Mat.

Process Piping

Instrumentation and Control

Insulation

Legal ExpensesContractor's FeeConstruction ExpensesCommissioning and StartupConstruction Management

Equipment Installation

Engineering and Supervision

Table 13. The overall cost estimation for the project

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When doing estimations, it should be noted that several assumptions are made in order to arrive at desired values. In particular, we would like to highlight that we minimized the need to consider Land and Site Preparation because we believe that it is a better alternative to rent out a space in an existing processing facility in Richmond than it is to buy land in the city instead (we are assuming that there is a facility already in place which would be able to rent us about 5000 square feet for about $10,000 per month). Not only is renting a less expensive option, it is also a safer investment as it minimizes potential lost funds in the future. Namely, should the plant experience any sort of technical difficulties or not be able to startup exactly as planned by the engineering phase, the plant can be shut down after however many months, losing only a few $10,000 worth of payments to rent. We have assigned Land and Site Preparation a 0% because we still believe there will be some site preparation to consider when renting.

Referring to table 16 below, the operating cost for this venture was calculated to be about $5.68 million per annum. This value is made in bulk of the cost of the cake (approximately $1.6 million based on 2,000,000 kilograms of cement being used and bought at $0.80 per kilogram as per the current market), and the cost of electricity (assuming a cost of $0.06 per kilowatt-hour in British Columbia). Other key items that were taken into consideration for the operating cost include (but are not limited to) the cost of reagents and consumables, maintenance, operating labor, and rental fees. Based on the total operating cost and the fact that we utilize about 2,000,000 kilograms of cake per year, we have determined that we will be able to safely purchase the cakes from HudBay at about $2.84 per kilogram of cake. The revenue we expect to obtain per kilogram of cake is about $4.11. Thus, per annum, we expect to have a revenue of about $8.2 million.

Price of Copper $6.11 /kg

Price of ZnCO3 $0.75 /kg

Cement Processed 8000 kg/day

Cement Processed 2000000 kg/year

Cu recovered 65.3 per 100 kg cake

ZnCO3 recovered 15.98127121 per 100 kg cakeTable 14. Calculated cost for the copper and zinc carbonate recovery

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REVENUE

Cu ZnCo3 TOTAL

Per 100 kg cake $398.77 $11.99 $410.76

Per day $31,901.92 $958.88 $32,860.80

Per year $7,975,480.80 $239,719.07 $8,215,199.87

Revenue per kg cake $4.11Table 15. Calculated revenue of cake per kilogram

Cost of Cement $0.80 /kg

ITEM% of CAKE

COST COSTDirect Production CostsCement cake Calculated $1,601,453.64Reagents Calculated $34,000.00Operating labor 38.35% $614,193.87Direct supervisory 7.67% $122,838.77Utilities 10.00% $160,145.36Electricity Calculated $1,452,758.40Maintenance 4.40% $70,463.96Supplies 0.67% $10,650.16Disposal Calculated $189,800.00Fixed ChargesInsurance 2.21% $35,466.08Local Taxes 4.42% $70,784.25Rental 7.75% $120,000.00General expensesAdministration 7.57% $121,184.38Distribution and selling 30.73% $492,078.89Research and development 15.36% $245,987.75Financing (interest) 20.88% $334,394.44Annual total operating cost $5,676,199.94Total operating cost per kg cake $2.84

Table 16. Calculated total operating cost per annum and cost of cake per kilogram

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YearTotal

Capital Investment

RevenueOperating

CostsGross Profit

Depreciation

Taxable Income Tax Return

Total Cash

0 $9.80 -$9.80 -$9.801 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 -$8.212 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 -$6.613 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 -$5.024 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 -$3.425 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 -$1.826 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 -$0.237 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 $1.378 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 $2.979 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 $4.56

10 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 $6.1611 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 $7.7512 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60 $9.35

13 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60$10.9

5

14 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60$12.5

4

15 $8.22 $5.68 $2.54 $0.65 $1.89 $0.94 $1.60$14.1

4TOTA

L $23.94IRR 14.0%

Table 17. Calculated IRR for the HudBay cake processing venture (dollar values reported in millions)

To further the economic analysis of this venture, we calculated an Internal Rate of Return (IRR) which is used to evaluate the appeal of project and this is tabulated in Table 17 shown above; in general, the higher a project’s IRR, the more desirable it is to undertake the project. For our endeavor, we have found the IRR to be 14%; we believe that this an attractive enough rate of return for our client to consider going through with the plan.

6.2 Environmental ImpactBy processing the cakes provided to us by HudBay, we produce a number of items. As mentioned prior, we are obtaining revenue from the copper and zinc carbonate we are recovering from the cake; another two products we are producing, however, are cadmium and lead. These two products raise particular concerns with regards to the environmental impact of our project as they are deleterious substances that need to be handled properly as to not incur undesirable outcomes.

Cadmium is obtained in significant amounts via processing and if it is relatively pure, we could potentially sell this cadmium to industries such as those of which pertain to the manufacturing of pigments, batteries, polymer stabilizers, coatings, and alloys (to

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name a few) and thus, it would serve to be another source revenue for the venture. However, to be conservative in the economic analysis, we have assumed that the cadmium we produce is not pure enough for sale and thus, we would have to pay a fee to impound it. In British Columbia, the Ministry of Environment is responsible for regulating and authorizing the quantity and quality of any discharge of waste to the environment. The cost to deport toxic substances is highly dependent on the jurisdiction it is being done in; for BC, we have assumed that it costs about $0.50 per pound (or $1.10 per kilogram) to impound toxic waste products. Thus, for the cadmium we produce, we can expect disposal costs to amount to about $101,600 per year. Similarly, lead is another heavy metal that does not have a comparably attractive market and for which we would have to pay a fee to deport. The fee to impound lead is about $88,200 per annum. In total, it would cost about $189,800 per year to dispose of the wastes created by our facility; this has been factored into the operating cost of the project.

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7.0 Recommendations

The purpose of this project was to determine feasibility of building and operating an ore residue processing plant in Richmond, BC. The plant recovers valuable metals such as copper and zinc by processing zinc plant residues. Also, hazardous wastes have to be safely disposed as the residues contain significant amount of lead and cadmium. The feasibility of plant operation was evaluated based on economics and environmental impact.

We have looked at three possible models for copper cake processing and performed economical analysis. The aim was to select feasible hydrometallurgical process for extraction of copper and zinc. Model 1 and 2 starts with oxidative acid leach using sulfuric acid whereas model 3 was leached by using ammonia. Copper is extracted from leach solution in all models and model 2 and 3 additionally precipitate copper. Zinc is recovered by solvent extraction in model 2 and zinc precipitation in model 3. The preliminary economic analysis was conducted assuming that reagents such as ammonia and sulfuric acid are 100% recycled. Our preliminary economic analysis indicated that the model three is the most preferable model for copper cement processing with highest return and we performed further detailed analysis for model 3.

Detailed economic analysis was conducted for model 3 based on capital and operating cost. Main factors to focus on the economics of processing were cost of building and managing the facility and the value of recovered products after processing. Capital cost was determined including specific equipment size based on flow rates. For more effective and inexpensive method, we considered EMEW cells for copper electrowinning. For conservative economics analysis, we did not consider lead and cadmium as additional source of at this stage.

Based on our analysis, building and operating a plant is recommended using ammonia leaching. Although ammonia leaching is a relatively new process, we concluded that ammonia leaching is the most profitable solution for copper cake processing. Considering current market price of copper cake of $0.80/kg, the plant will derive satisfying revenue with about 14% internal rate of return by recovering precious metals while safely disposing hazardous lead and cadmium. For further improvement, we suggest future analysis on cobalt cake and alternative way of disposing lead and cadmium for another source of revenue.

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8.0 Project Planning

After a thorough discussion, specific sections were distributed to each team member. In order to carry out a preliminary mass and heat balance for the processing of copper cement cake, appropriate reagents and compounds are required and this will be based on the designed flow sheet. The best approach for the project is to have each design task assigned within a timeframe so that each step of the project can be completed accordingly and correctly. Most of the time, organization of the project took place during the project review meetings held every Wednesday alongside Professor Edouard Asselin. The progress of each team member were discussed and shared through Facebook and email.

Flow sheet development was the first crucial task that had to be completed accurately. Three different design options were examined by all team members and were constructed properly by Dishoo. He analyzed all three of the flow sheets and carefully designed it based on its metal recovery. The preliminary mass and heat balances were carried out by Nadeeshika and Estelle. They had to take into account of the specific reagents needed for each option and accurately choose the chemicals that will help maximize the metal recoveries. With this, they managed to generate an estimate of the metal recoveries as well as the amount of the different reagents involved.

Based on the preliminary economic analysis completed by Dishoo, design option 3 was selected and a detailed mass and heat balance were carried out by Harith and Taehyun. Harith and Taehyun completed mass balance sheet by mid-March but there were some issues with the mass balance and be critical on certain aspects including:

1. The solubility of solid in water

2. Solubility in liquid ammonia

3. The mass loss from solvent extraction

4. Electro-winning operation and how does the ammonia get recycled from electrowinning

5. Ammonia consumption due to recycling from zinc precipitation and electrowinning

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Design Tasks Team Members

Background Information All MembersFlowsheet Options All MembersPreliminary Mass and Heat Balances Nadeeshika and Estelle (Model 1 and 2)

Harith and Taehyun (Model 3)Preliminary Economic Analysis DishooDetailed Mass and Heat Balances Taehyun and HarithEquipment Sizing Dishoo (Sizing) and Jeff (drawing Equipment

Diagram)Cost Estimation Nadeeshika and Estelle

Design Report

They managed to determine the total amount of ammonia consumed the sum of copper extracted, and the total of cadmium and zinc recoveries. Sizing of the major equipment were finalized by Dishoo where the feed flow rates and volumes for each tank were determined by utilizing the mass balance. He had successfully calculated the suitable tank size for water and ammonia leaching, copper extraction and the recoveries of cadmium and iron. This leads to the final step of the project where an approximation of the capital and operating costs were done by Nadeeshika and Estelle. Other group members had also assisted to this socio-economic parts by checking if there is any error in calculation in order to make further recommendations in the future for this kind of hydrometallurgical plant utilizing ammonia leaching as the main basis to leach the metallic components of the concentrates. Taking the environmental and safety aspects of the plant into consideration, they managed to generate an acceptable purchase price for the copper cement. Below is a Gantt chart generated to illustrate the time spent on each design task up to the end of the project.

Introduction

Literature Review

Project Distribution

1. Flowsheet Development

Cadmium and Copper Recovery

Cadmium, Copper, and Zinc Recovery

Modified Copper, Cadmium and Zinc Carbonate Recovery

2. Mass and Heat Balances

Identify Reagent Requirements for Each Design Option

Perform Economic Analysis

3. Equipment Sizing

Anaylze Appropriate Tank Sizes and Feed Rates

Perform Flow Rate and Tank Volume Calculations

4. Cost Estimation

Compute Total Operating and Capital Costs

Establish Purchase Price for Copper Cement

5. Generate Final Report

Executive Summary

Problem Specification

Technical Review

Detailed Design

Socio-Economic Assessment of Design

6. Presentation Day

Powerpoint Preparation

Present

16-Jan 23-Jan 30-Jan 6-Feb 13-Feb 20-Feb 27-Feb 6-Mar 13-Mar 20-Mar 27-Mar

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9.0 References

1. Agency for Toxic Substance and Disease Registry (ATSDR). (1989). Public Health Statement for Cadmium. Atlanta, GA: U.S. Department of Health and Human Services

2. C.L. Pfalzgraff, “Do’s and don’ts of tankhouse design and operation,” in Copper Leaching, Solvent Extraction, and Electrowinning Technology, G.V. Jergensen II, ed., Society for Mining, Metallurgy and Exploration, pp. 217-221

3. Cognis Group, MCT Redbook. Solvent Extraction Reagents and Applications, 2007. Retrieved Feb.21/09 at:http://www.cognis.com/NR/rdonlyres/7A687186-A305-4969-B6EA-0CF869930714/0/MCT_Redbook_English.pdf

4. C.R. Merigold, Lix Reagent Solvent Extraction Plant Operating Manual, Cognis Corp., 2nd edn, 1996.Retrieved Feb. 18, 2009 at: http://www.cognis.com/NR/rdonlyres/97088921-AD31-461E-A120-9C2B791ACD26/0/lixsolve.pdf

5. Diankun Lu, Zhenan Jin, Lifang Shi, Ganfeng Tu, Feng Xie, Edouard Asselin (2014). A novel separation process fordetoxifying cadmium-containing residues from zinc purification plants. Elsevier Minerals Engineering

6. Electrometal Technologies Limited (2014). “Performance Characteristics of the EMEW Cell”. Retrieved from https://www.electrochem.org/dl/ma/203/pdfs/2376.pdf

7. Ettel, V.A. in Hydrometallurgy: Theory and Practice Course Notes Centre for Metallurgical and Process Engineering, UBC, p. X-7.

8. G.A. Kordosky, R.B. Suddert and M.J. Virnig, “Evolutionary development of solvent extraction reagents: real-life experiences,” in Copper Leaching, Solvent Extraction, and Electrowinning Technology, G.V. Jergensen II, ed., Society for Mining, Metallurgy and Exploration, pp. 259-271.

9. G. Kordosky, "Copper recovery using leach/solvent extraction/electrowinning technology: Forty years of innovation, 2.2 million tonnes of copper annually," Journal of South African Institute of Mining and Metallurgy, 2002, Nov.-Dec., pp. 445-450.

10. J. Jenkins, W.G. Davenport, B. Kennedy and T. Robinson, “Electrolytic copper -leach, solvent extraction and electrowinning world operating data,” Proceedings of the Copper 99 – Cobre 99 International Conference, S.K. Young, D.B. Dreisinger, R.P. Hackl and D.G. Dixon, eds., Phoenix AZ, Oct. 10-13, 1999, Vol. 4, pp. 493-567.

11. J.R. Spence and M.D. Soderstrom, “Practical aspects of copper solvent extraction from acidic leach liquors,” in Copper Leaching, Solvent Extraction, and Electrowinning Technology, G.V. Jergensen II, ed., Society for Mining,

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Metallurgy and Exploration, pp. 239-257. (b) A very slightly different version of this article can be accessed at (Feb. 21/09). Retrieved from http://www.cytec.com/specialty-chemicals/pdf/practical.pdf

12. Lu, D., Jin, Z., Shi, L., Tu, G., Xie, F., Asselin, E. (2014). A novel separation process for detoxifying cadmium-containing residues from zinc purification plants. Elsevier Minerals Engineering.

13. M.S. Prasad, V.P. Kenyen and D.N. Assar, “Development of SX-EW process for copper recovery-an overview,” Mineral Processing and Extractive Metallurgy Review, 1992, vol. 8, pp. 95-118.

14. Organization for Economic Co-operation and Development. (1995). Recycling of Copper, Lead, and Zinc Bearing Wastes. Paris, France.The Province of British Columbia. (2014). Transporting hazardous waste. Retrieved from http://www2.gov.bc.ca/gov/topic.page?id=75593F8559104D9E8C8ED75003A18301&title=Transporting%20Hazardous%20Waste

15. United Nations Environment Programme (UNEP). (2011). Study on the possible effects on human health and the environment in Asia and the Pacific of the trade of products containing lead, cadmium and mercury

16. W.G. Davenport, M. King, M. Schlesinger and A.K. Biswas, Extractive Metallurgy of Copper, 4th edn., Pergamon, 2002.

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